-------
'Chromium removal is accomplished by first oxidizing the chromium
with chlorine gas; electrochemically or potentially with SO--CL gas
mixtures, then precipitating the uichromate ion as lead chromate.
Oxidation has been shown to be effective in laboratory scale test
reactors. Large scale oxidation testwork using chlorine and an
electrochemical reactor have been performed successfully.
A recycle system for stripping the oxidized chromium from the leach
solution has been operated succes:fully: the solution is exposed :o
lead sulfate in an agitated reactor; lead chromate precipitates; the
lead chromate product is crystalline and dense and settles rapidly;
the solution essentially free of lead chromate solid is pumped from
the solids for further treatment for nickel removal; the lead
chromate is redissolved in sulfuric acid to form a concentrated
chromic acid solution and lead sulfate; the lead sulfate solid is
separated from the chromic acid and is recycled to the lead chrcmate
precipitation reactor.
'Nickel can be removed by sulfide precipitation. The reaction is
rapid and near quantitative. The pH is maintained in the range 4-5
so hydrogen sulfide is not released. The solid product is reaoily
filterable. Quantitative removal of nickel is not necessary because
practically all the final solution can be recycled to the
leach-jarosite precipitation unit operation. Therefore, the
addition of a deficiency of sulfide (less than the stoichiometrlc
requirement for complete nickel removal) is desirable so that all
the added sulfide ions are consumed. Then when the solution Is
recycled to the acid leach step hydrogen sulfide gas will not be
formed. Other alternative nickel recovery unit operations are
discussed later. An attractive alternative is the production of
nickel oxide (Section 6.4).
2.3. ECONOMIC ANALYSIS
An "order of magnitude" estimate has been performed on the flowsheet
presented in Figure 2.1 and expanded in Figure 6.7. The calculated return on
Investment (ROI) based on the "order of magnitjde" estimate is normally
considered to be within *30%^49i50).
Definitions and cost estimation factors are taken primarily from Mular
"Mineral Processing Equipment Cost and Preliminary Capital Cost Estimations",
and Wood, Chapter 29.1, "Cost of Equipment", and Pratt, Chapter 29.2, "Cost of
Process", in the Solvent Extraction Handbook. A summary of the assumptions
made and the detailed calculations for treating fifty tons of sludge per day
are presented in Section 6.4 and Appendix 8.15. The major assumptions include:
16
-------
the land and buildings are available; a credit of one dollar per gallon of
sludge is allowed; and the tax rate is fifty percent.
The results of the calculations are tabulated for each unit operation in
Tables 2.1 and 2.2. The first order estimate for the return on investment is
41 +. 12%.
The largest cost unit operation is recovery of chromium; oxidation is by
far the costliest step in recovering useful chromium compounds. There is
potentially a new low cost oxidation process now being commercialized for
cyanide destruction. The solution oxidizing potential has been shown to be
high enough to oxidize nickelous ions in solution to form nickelic hydroxide
solid. That level of solution oxidizing power will certainly oxidize chromium
(Cr*3 —> Cr*6). The process (S02 + 02) is described in Section 6.4. Such a
process would not only be less capital intensive but the energy savings would
be great. A cost comparison between the flowsheet presented in Figure 2.1 and
the flowsheet modified for SO.-Og oxidation is presented in Table 2.3. The
difference in the ROI is significant; 41% for the flowsheet presented in Figure
2.1 and 691 for the S02-02 modified flowsheet.
•
•
Another potential alternate treatment process is solvent extraction and
electrowinning of nickel, precipitation of chromium hydroxide, and production
of chromium oxide (discussed in Section 6.4). A cost comparison between the
flowsheet presented in Figure 2.1 and the modified flowsheet is presented in
Table 2.4. The difference in the ROI is significant; 411 for the flowsheet
presented in Figure 2.1 and 67% for the modified flowsheet.
The detailed cost analyses results presented in Section 6.4 and 8.15 show
good potential for an excellent return on investment. Even if a credit is not
taken for disposal, two of the modified flowsheets show an income sufficient to
offset tne cost of the treatment process. It is recommended that further
consideration be given to the economic consequences of variations in the chosen
unit operations.
17
-------
TABU" 2.1 PROCESS COST: FIRST ORDER ESTIMATE
Unit Operation . COST (S)
Factored Capital Annual tied Capital Operation Cost Total Cost
Cost Estimate Cost Per Year Per Year
I. Leach, jarosHe
precipitation 430.800 • 119.500 223.500 343.000
2. Jarosite storage 390.500 108.200 25.400 133.600
3. Copper solvent
extraction, electro-
winning 336.100 93.100 205.403 299.000
*
4. Zinc, residual iron
solvent extraction,
zinc sulfate crystal-
ization 661,POO 183.300 269,700 453.000
5. Chromium oxld..
chromic acid pro-
6.
duct ion
Nickel recovery
TOTAL COST
1
3
.818.
231.
,868.
200
600
800
503.
64.
1,0?)
600
200
.900
407.
230.
700
000
1.362.200
911.
294.
300
200
2.434,100
See Section 6.4 for details.
-------
TABLE 2.2 PROCCSS COST SUMMARY: FIRST ORDER ESTIMATE
Unit Operation COST (i)
Factored Capital Operation Cost lotal Cost Potential value
Cost/Vr P I2X Per Yr Per Vr of Product«/lb)
1. Leach, jarosite
precipitation 227.700 4.0* 248.900 4.4* 476.600 8.4*
2. Copper SX. EW 93.100 25.0 205.900 S5.2 299.000 BO.2 60
3. Zinc, residual
Iron SX, zinc
sulfate cryit. 183.300 17.4 269.700 25.7 453.000 43.0 20
4. Chromium oxid.,
chronic acid
production 503.600 66.1 407.700 53.S 911.300 119.6 118
5. Nickel Recovery 64.200 10.9 230.000 39.0 294.200 49.9 172
• per pound of residue solids.
See Section 6.4 for details.
-------
TABLE 2.3. COUPARSION OF FIRST ORDER COST ESTIMATES BETWEEN FLOWSHEETS FOR ELECTRO-
CHEMICAL OXIDATION AND S02 - Oj OXIDATION OF CHROMIUM
Flowsheet COST (I)
FCC FCAC Operating Total Product Value*
Cost/yr Cost/yr
Electrochemical 3,868.800 1,071,900 1.362,?00 ?,434.100 5.643.400
(UbleZ.I)
Modified 2.862.900 793,300 1.209,100 2,002.400 5.885,800
R.O.I. •[(S.R85.800 - 2.002.400)/2.862.900 ](0.50)(100)
• 69 t 20 t
• Same products In both flowsheets except for nickel- HIS In Table 2.1. NtO It modified
flowsheet.
See Section 6.4 for details.
-------
TABLE 2.4. COMPASSION OF FIRST ORDER COST ESTIMATES BETWEEN FlbKSHEETS FOR
OXIDATION AND NICKEL SOLVENT EXTRACTION AND RECOVERY.
Flowsheet FCC FCAC
Electrochemical 3,866,800 1,071.900
Modified 2.477,300 824.900
Operating Cost
Per Year
1.36?. 200
1.175.500
Total Cost
Per Year
2.434.100
2.000.900
ELECTROCHEMICAL
Product Value*
5,643,400
5.977.100
ROI °[(5.977.100 - 2.000.900) / 2.977.3001(0.5)1100)
• 67 ! 20 8
• Sane products in both flowsheets except for nickel (nickel in modified flowsheet) and
Chromium (chromium oxide In modified flowsheet).
See Section 8.15 for details.
-------
SECTION 3
RECOMMENDATIONS
The treatment of hydroxide sludge materials for metal value recovery by a ser-
ies of conventional extractive metallurgical unit operations has been demon-
strated. The treatment sequence is selective and effective for recovering
copper, zinc, chromium, and nickel. Iron, calcium and aluminum can be
extracted from the leach solution and rejected from the system.
The highest cost unit operation in the treatment sequence is oxidation of
chromium. Alternatives have been suggested; oxidation by SOg/Og gas mixtures
or solvent extraction of nickel, precipitation of chrom1um(3) hydroxide with
subsequent calcining to chromium CKide. Both of these alternatives to the
original flowsheet appear to offer a great savings in cost. The alternatives
are, however, not presently commercially proven processes. Further research
and development studies are necessary to insure applicability to the present
system. Specifically the needed research includes:
1. A study of the possibility of oxidizing chromium In a chromium-
nickel bearing solution by SO./O- mixtures. The use of SO./0. is
presently commercially used by InCO to destroy cyanide in waste
leach solutions. The oxidizing potential that can be developed
by the SO./O. has been shown to be sufficient to oxidize
nickel(+27 to nickel(+3). Therefore, the application of SO./O-
to a chromium-nickel solution appears to have the potential for
oxidizing both chromium(+3) and nickel(+2). The envisioned
treatment would be carried out at a low solution acidity, i.e.,
pH 8. The chromium and nickel would both exist as hydroxides.
As the oxidization progressed chromium(+6), as chromate, would be
soluble; nickel(+2) would be oxidized to nickel(*3) hydroxide and
remain as a solid. A solid/liquid separation would then be used
to separate the chromium from the nickel. The chromium(+6)
solution could be treated as suggested in the previous
flowsheets. The separated solid nickel hydroxide could be
calcined to nickel oxide.
2. A study to determine the possibility of separating nickel from
chromium by solvent extraction using either mixtures of
22 .
-------
D-EHPA-EHO or D-EHPA-LIX63. Both have been shown previously by
other Investigators to extract nickel from acidic solutions.
Investigations reported in the present study show that the
extraction is selective toward nickel, i.e.. in a nickel-chromium
bearing solution, nickel is extracted but chromium(+3) is not.
Therefore, an envisioned treatment would include selective
removal of nicfcel(+2) with subsequent recovery from solution as
nickel by electrowinning; followed by precipitation of
cnromium(+3) hydroxide at a pH of 3-4.5; and then solid/liquid
separation with subsequent conversion of chromium hydroxide to
chromic oxide by calcining.
It Is recommended that a detailed cost analysis be performed on the
proposed flowsheets and the potential alternate unit operations. The first
order cost estimates presented in Section 6.4 indicate that sludge treatment
may be economically attractive. These estimates, therefore, should now be
followed by detailed cost projections.
23
-------
SECTION 4
MATERIALS AND METHODS
4.1. SLUDGE CHARACTERIZATION
4.1.'l. Starting Sludge Material
Experimental analytical procedures and sample preparation techniques are
presented in Appendix 8.1. For the most part. Induction Coupled Plasma
Spectrophotometry (ICP) was used for determination of elemental concentrations
in solutions.
4.1.1.1. Phase I Material
Sludge materials were obtained from three different industrial sources.
The material was packed in fifty-five gallon barrels by the producer and
shipped to Butte, Montana. The sludges were. In most cases, mixed metal
hydroxide materials (Table 4.1). A portion of the supplied material was
electroplating cell bottom sludge rather than precipitated hydroxide sludge,
e.g., 6, 7, and 8. The solids content of all the sludges ranged approximately
20-35 weight percent, e.g.. Table 4.2.
Even though the sludge materials were only 20-35 percent solids they could
be handled like solids, i.e., they could be broken into smaller pieces without
release of free water. The material could be broken up into small pellet-like
chunks (approximately one-eighth inch diameter) by use of a laboratory hand
mixer.
X-ray diffraction patterns of dried sludge showed the material to be
amorphous :as is typical of precipitated hydroxides.
24
-------
TABLE 4.1. NIXED METAL SLUDGE COMPOSITION OF AS-RECEIVED SLUDGE
IM
cn
Sample No. Sludge Source
Barrel 1
544 A
544 B
545 A
54S B
546 A
546 B
Barrel 2
927
928
929
975
976
Barrel 5
227
228
229
986
987
988
989
*
Composition U) In Solids
Cu
8.06
7.87
7.56
7.69
7.63
8.24
5.66
5.61
5.84
5.69
5.11
2.41
2.41
2.48
2.46
2.26
2.54
2.41
Fe
18.21
17.70
19.16
18.71
17.19
18.65
15.90
15.75
16.58
15.17
14.32
11.33
11.88
11.65
12.68
12.18
13.34
13.11
In
11.73
11.46
10.98
11.11
11.58
11.94
10.76
10.67
11.18
10.15
9.4C
8.40
3.45
8.7*
8.72
4.18
3.84
8.66
Cr
.20
.19
.11
.14
.13
.23
1.23
1.23
1.29
1.03
0.98
.36
.35
.35
.10
.08
.18
.16
HI
5.59
5.55
5.30
5.44
5.41
5.86
6.11
6.07
6.31
4.16
3.86
5.08
4.80
5.08
3.69
3.52
3.84
3.65
Cd
0.74
0.72
0.70
0.71
0.71
0.77
0.66
0.6"
O.o,
0.52
0.48
0.39
0.41
0.40
0.29
0.23
0.25
0.30
Al
2.92
2.78
2.68
2.75
2.71
2.94
4.64
4.60
4.83
3.94
3 74
4.05
4.15
4.55
5.23
4.87
5.74
5.63
Ca
1.00
1.19
1.01
1.00
2.47
1.08
1.41
1.38
1.46
0.76
n.90
.08
.00
.10
.04
.07
.11
.08
P
.13
.51
.29
.44
.49
.98
2.79
2.93
3.13
4.30
4.39
• • ••
• • . •
....
2.03
2.66
2.40
2.46
Barrel 6
1036
0.12
-------
TABLE 4.1. CONTINUED
Sample No. Sludge Source
Composition (t) in Solids
i\>
1037
1038
1039
1223
1224
Barrel 7
Barrel 8
Barrel 9
Barrel 10
1040
1041
1222
2135
Barrel 11
Barrel 12
Barrel 13
Barrel 14
Cu Fe Zn Cr Nt Cd A) Ca P
0.16 0.004 0.18 0.54 30.31
-------
TABLE 4.1. CONTINUED
Sample No.
1700
1701
1702
1703
1820
1821
1822
1225
2136
2137
2138
2!39
Sludqe Source
Cu
Barrel 14 (cont.)
2.29
2.19
2.25
2.05
2.10
2.19
2.14
Barrel 15
1.59
Barrel 16
1.70
Barrel 17
4.00
Barrel IB
' * 6.78
Barrel 19
1.91
Composition (X) 1n Solids
Fe
17.42
18.49
18.70
18.26
1C. 40
19.55
20.05
17.52
15 68
15.20
17.19
18.68
Zn
11.65
10.32
10.41
10.57
9.24
9.48
9.16
9.64
16.92
10.53
6.81
13.31
Cr
.14
.05
.09
.04
.03
.07
.10
1.75
1.51
4.90
7.13
2.67
N1
9.40
8.76
8.82
8.35
8.70
8.56
7.12
10.20
3.83
3.89
2.31
4.56
Cd
0.39
0.42
0.43
0.46
0.40
0.41
0.47
0.32
0.37
0.16
0.01
0.15
Al
•^•IV^^BHVV
1.97
2.25
2.30
2.27
2.18
2.33
2.48
<0. L.
•3.70
2.62
2.44
2.28
Ca
Calcium
Channel
not
Operative
I
1
ff
5.94
1.38
0.98
0.23
0.97
_P
3.19
2.87
3.03
2.85
2.19
2.22
2.31
1.26
1.26
2.23
3.19
1.67
Sludge solid content varied from approximately 20-30 weight percent solids.
-------
TABLE 4.2. MOISTURE CONTENT OF AS-RECEIVED MIXED METAL SLUDGE
Sludge Source
Barrel fl
Barrel 12
Barrel 15
Barrel 16
Barrel 17
Barrel 18
Barrel 19
Barrel 110
Barrel 111
Barrel 112
Barrel 113
X HjO
76.34
77.69
76.20
77.01
77.62
7/.BS
76.90
77.67
76.39
76.59
76.13
76.44
77.46
76.23
82.45
83.89
79.41
82.25
76.69
79.50
X Solids
23.19
76.81
22.47
77.53
23.61
76.39
22.54
23.77
17.55
16.11
20.59
17.75
23.31
20.50
-------
TABLE 4.Z. CONTINUED
ro
«o
Sludge Source
Barrel 114
'Barrel 115
Barrel 116
Barrel 117
Barrel 118
Barrel 119
I H20
81.34
82.13
81.43
81.80
82.75
82.45
81.18
76.69
70.65
74.19
69.47
77.48
I Solids
81.87 18.13
23.31
29.35
25.81
30.53
22.52
-------
A water leach of starling sludge material showed very little redissolution
of metal values, e.g., sludge from barrel number eight (Sample No. 1355) showed
very little metal dissolution: 0.6% Cr, 1.3% Fe, 1.4% Ni, 1.9% Cu, 2.2r. Al
•
(leach conditions: 10% solids, 0.5 hr., ambient temperature).
• •
4.1.1.2. Phase II Material
The material used in the Phase II study was obtained from the local
California company where the test assembly was located. The required test
material was obtained as needed from current daily sludge production. Example
analyses an> presented in Table 4.3. The sludge was primarily a high
chromium-high nickel-low iron material. The solid content varied between 16-30
percent. In some cases, the sludge material was doped with copper and zinc
sulfate for testwork requiring solutions containing iron, nickel, chromium,
copper and zinc.
4.1.2. Methods of Analysis
A detailed sum..-.ary of sample preparation and analytical procedure used to
chemically, characterize the sludge materials is presented in Appendix Section
8.1. The sample dissolution procedure used was a perchlorate fuming technique:
the aqueous solution analytical technique used was atomic absorption and
induction coupled plasma spectrophotometry. Montana Tech Foundation was
supplied with a set of solutions by EPA to verify the laboratories' analytical
capabilities. The E?A solution analytical verification results are reported in
Appendix Table 8.1. and discussed in Section 8.1. All aqueous leach solutions,
raffinates and organic analyses were performed using induction coupled plasma
spectrophotometry.
4.2. REAGENTS
Chemical reagents used throughout the study were either technical or
reagent grade. They included: acids; bases; solid compounds such as lead
sulfate, sodium sulfide. Tap water was used in the large scale testwork;
30
-------
TABLE 4.3. MIXED METAL SLUDGE COMPOSITION FOR CAMARILLO SLUDGE MATERIAL
Sample No.
3177
3183
3184
3291
3301
3332
Solid Content
16.5
25.0
25.0
26.5
28.7
26.7
U)
Cu
0.1
0.1
0.1
0.1
0.1
0.1
Composition (%) \n Solids
Fe
5.3
5.8
4.8
4.5
4.7
2.4
Zn
0.5
0.3
0.3
0.4
0.1
0.3
Cr
5.0
13.3
11.4
9.1
11.7
10.5
N1 Al Pb Ca P
32.0 <0.1
17.9 <0.l <0.1
19.7 <0.1 <0.1
34.4 <0.1 <0.1 0.4 -
23.3 <0.1 <0.1 0.8
26.1 <0.1 <0.1 0.9 -
-------
deionized water was used in sma'.l scale kettle testwork and in all reagent
dissolution and dilution procedures.
Solvent extraction reagents were supplied by vendors and were the same
reagents supplied to their commercial customers. The reagents were sometimes
donated to the project and at other times were purchased. The reagents
Included: LIX-64N, LIX-622, LIX-70.(Henkel Corporation); D2EHPA and Alamine
336 (Mobil Corporation); ACORGA 5IOC (ACORGA Corporation); ONNSA and XB-1 (King
Industries). These reagents were diluted to the desired strength by use of
KERMAC 470B and KERMAC 510B Kerosene (Kerr-McGee Corporation).
Ion exchange resins were supplied by Rohm and Haas and were the sane
resins supplied to their industrial customers. Those resins used in thie study
included: a weakly basic anion cation exchanger IRA-94; a strongly basic anion
exchanger IRA 900; and a strongly acid cation exchanger IRA 200.
32
-------
SECTION 5
EXPERIMENTAL PROCEDURES
5.1. LEACH AND PRECIPITATION STUDIES
The leach and precipitation studies were Initially conducted in one liter
thermostated reaction kettles. A typical set-up is presented In Figure 5.1.
The reaction kettles allowed testwork to be controlled over a wide range of
experimental conditions. A two level factorial design matrix was utilized In
order to minimize the number of experimental tests necessary to establish
appropriate experimental conditions for the larger scale testwork.
Experimental conditions investigated for the leach and jarosite precipitate
testwork includad: reaction temperature; reaction time; acid and reagent
concentration; solution Eh; agitation rate; and solid/liquid ratio.
The conditions for each Individual experimental study were based on the
set of conditions specified in the design matrix table. For example, the
experimental leach study procedure Included: selecting and blending a starting
sludge sample; splitting a sample for determination of moisture content;
splitting a sample for determination of elemental content; weighing the sludge
sample and placing in the reaction kettle; setting the experimental temperature
(thermostated water bath); Initiating the study by addition of concentrated
sulfuric acid; diluting the sample to the desired volume; setting the agitation
rate; adjusting pH; sampling the solution as a function of time for analysis;
running the test for designated time; removing the reaction kettle from the
bath vessel; separating the solid from liquid by vacuum filtration and sampling
the solution and sometimes the solid for analysis. Based on the results from a
series of design matrix tests and further optimization testwork, a standard
leach procedure was adopted; i.e., one-half hour; temperature 40-S5°C; add
concentration, equivalent to a weight of 100% of the solid content of the
33
-------
Figure 5.1. Laboratory leach system.
34
-------
sludge; a sludge/liquid ratio of 0.8; and an agitation rate to completely
suspend all particles in the solution phase.
Jarosite precipitation studies folloxed the leach studies. Testwork was
performed to determine the appropriate conditions for jarosite precipitation of
iron from the leach solutions and the appropriate conditions for jarosite
precipitation in the presence of leach residue solids (designated in-situ
precipitation). The procedure was to first leach the sludge material under
standard conditions then to either filter the solids from the solution or to
leave the solids In the solution and to then adjust the conditions to permit
jarosite to form. The results of these studies are presented in Section 6.2.2.
and Appendix 8.3.1.
The small-scale testwork was followed by leach and jarosite precipitation
experiments in a ninety-liter poljrprooylene reaction vessel; followed by
testwork in a full-scale 270-liter polypropylene reaction vessel. (The results
of these experiments are presented in Section 6 and Appendices 8.3.1 and 8.13.
The experimental procedure for the large-scale testwork was similar to the
laboratory testwork. The experimental conditions for the large-scale testwork
were based on the best small scale results. A schematic drawing of the
leach-jaroslte reaction vessel is presented in Figure 5.2; a pictoral depiction
is presented in Section 8.14.
The experimental procedure for the large scale high iron bearing sludge
testwork included: sludge blending and sampling; feeding into the 270 liter
reaction vessel; adding concentrated sulfuric acid slowly to break up the solid
chunky material; diluting to the desired volume (this process raised the
temperature to 50-60°C); placing a heavy duty stainless steel agitator in the
reaction chamber to suspend the solids In solution; reacting for one-naif hour;
raising the temperature to approximately 90°C (by two 6,000 watt quartz
immersion heaters) adjusting solution pH conditions to 2.2-2.6 using KOH;
adding K2S04 so that tne stoichiometry and reaction conditions were appropriate
for jarosite precipitation; reacting for 4-6 hours (pH periodically adjusted);
sampling hourly to determine the iron content of the solution; adding dropwise
(at about 1,000 cc/hr for the last two hours of the test) hydrogen peroxide to
35
-------
Mr Driven Aglmor
6000 watt Immersion
Heater and Guard
(Two Required)
Plastic Top .
Steel Shell
Figure 5.2. Le«ch-jarosite reaction vessel (Cross section view)
36
-------
oxidize the ferrous iron; pumping the solution to n storage tank for
solid/liquid separation by settling (required about one-half hour for complete
settling); pumping the solution from the settling tank to a feed tank for the
following SX unit operations; and pumping the jarosite loaded slurry, about 401
solids, to the LASTA filter press (described in Section 3.5) for final
solid/liquid separation. The filter cake was sampled to determine moisture
content and to determine If the solids would pass the EP Toxicity Test.
The experimental procedure for the large scale low iron bearing sludge
excluded the jarosite precipitation unit operation. The sludge material was
blended; fed Into the reaction vessel; sulfuric acid solution was added and the
leach reaction was initiated and conducted for one-half hour. The resulting
slurry was pumped to the LASTA prers and filtered using a filter aid.
5.2. SOLVENT EXTRACTION
Studies were conducted to investigate the potential application of solvent
extraction (SX) to selectively extract and recover copper, zinc, iron, and
nickel. The experimental methodology consisted of first conducting batch shake
tests on a small scale (125-250 cc) in separatory funnels. These preliminary
experiments were followed by continuous testing in a Bell Engineering 600 cc
mixer-settler test rack; followed by full-scale continuous testing in a Relster
one-gallon mixer-settler test rack.
The hand shake tests were performed to establish: the influence of
reagent selectivity for a particular element; the Influence of aqueous phase
pH, temperature, time, diluent concentration, and reagent concentration, on
chemical specie exchange and phase separation between the organic and aqueous
phase during extraction and during stripping operations. The shake tests
provided a means for selecting appropriate conditions under which to start the
continuous testwork.
The small scale test rack consisted of ten 600 cc mixing chambers and ten
•
600 cc settling chambers. A combination of one to ten cells could be assembled
so the counter-current flow, and contact and settling of the organic and
37
-------
aqueous phases were controllable. Solution flow rates (to 50 cc/minute)
between mixers and settlers and the organic-aqueous interface positions were
controllable. Therefore, retention time and organic/aqueous phase ratio were
controllable.
The larger scale test reCK consisted of ten one-gallon mixing chambers and
ten one-gallon settling charters. Solution flow rates were controllable up tc
500 cc/minute. Details of the solvent extraction system are presented
schematically In Figures r>.3 and 5.4.
Two large scale test racks were available for the project. Individual
cells were connected in a variety of arrangements to study both copper and zinc
extraction from the aqueous phase and to study stripping characteristics of the
metal values from the organic phases.
5.2.1. Copper Solvent Extraction
5.2.1.1. Separatory Funnel Shake Test
The small scale separatory funnel (125 and 250 cc) shake tests were used
to Investigate the applicability of a specific extracting reagent to the mixed
metal aqueous solutions. The experimental procedure used In the testwork
followed the sequence; the pH of the aqveous phase was adjusted to the desired
value; an organic phase was prepared containing a specific extracting agent
dissolved In a kerosene solvent; the two phases were added to the separatory
funnel'In the desired organic to aqueous ratio (0/A); the separatory funnel was
stoppered and agitated for a specified time; the agitated mixture was allowed
to separate into two distinct phases and each phase was sampled for analysis;
the pH of the aqueous phase was measured to establish the equilibrium pH.
5.2.1.2. Large Scale Test
The large scale testwork was performed in the Reister testrack.
Preliminary continuous tests were performed in the smaller Beli Engineering
testrack to establish proper mixing and settling residence time and to
determine if muck or crud formation would be a problem.
38
-------
I _ I _
END VIEW
SX CEL
^•^IH
Figure 5.3. Relster one-gallon oixer-settier system: Testrack details.
AGITATOR
SPFEO CONTROL
-------
SIDE VIEW
ORGANIC
OUTLET
TOP VIEW
Figure 5.4. Reliter one-gallon mixer-settler system: Individual
-------
The procedure used in the testwork followed the sequence: a decision was
made on the number of extraction stages (one and two Investigated) and
stripping stages (usually two stages); the stages were connected so that a
countercurrent aqueous-organic flow pattern was established (Figure 5.5); the
cells were loaded with the proper organic tc aqueous ratio (0/A); solution flow
was Initiated at desired flowrate (up to 500 cc/min.); samples of aqueous
raffinate and strip acid into and out of the system were pulled as a function
of time; pH of the raffinate was monitored.
5.2.1.3. Organic Degradation Testwork
Extended exposure testwork was conducted to determine if the organic phase
showed extensive degradation and deterioration with continued use. The Bell
Engineering SX rack was used for this testwork. Three stages of extraction and
two stages of strip were investigated. The test conditions were similar (but
for extended times) to the large scale testwork; 50 cc/min. (250 cc/m1n. large
scale testrack); 0/A » 1 for both organic loading and stripping; pH • 1.75;
temperature, 50-55°C; and strip acid 200 gpl H-SO^.
The procedure used In the testwork was to expose a fixed volume of organic
( 3 liters) to a large quantity of copper bearing leach solution. The organic
solution was repeatedly exposed to copper loading and stripping. The
effectiveness of the organic phase was determined by closely monitoring the
element concentrations in the raffinate solution and by sampling the organic
phase after approximately every forty liters of aqueous contact. The organic
sample was stripped twice with 200 gpl suIfuric acid then exposed to a standard
leach solution (two contacts). The effectiveness of the organic extractant was
determined by its ability to remove Cu selectively from the standard solution.
5.2.2. Zinc Solvent Extraction
5.2.2.1. Separatory Funnel Shake Test
The small scale testwork was conducted using the same procedure outlined
In Section 5.-2.1.1. D^EHPA was the only extractant investigated for zinc
41
-------
ro
Loaded Organic
(Pimp Required)
Extraction 1
Stage One |
6
o
SETTLER
MIXER
(
}
Xj.
i
Aqutout
• .
•>^.
r»«!7?*^.
Extraction
Stage Two
— «o*-"
MIXER
SETTLER
9
••o \
•""•^""^ .
•
Strip
Stage Two
.JO
o
SETTLER
MIXER
QsS — -
4r^"*-
I
<
IStrip
(Stage One
-J"'P Held
^^^^^^^^^^^^^^••i
lt'i» RtCjel.
i^jr-^
i
j
I
i
MIXER
SETTLER
(j
-o
i i
;>
r
Aqueous Feed Aqueous Product Return Strip Acid To Cu E.W. or
(fu.p Rtquirtd) (n.ffin.u) from Cu E.H. or Crystallization
Crystallization
(Puap Rtquirtd)
Figure S.S. Copper solvent extraction flow pattern.
•
-------
extraction. Phase separation proved to be a problem for high Iron bearing
solutions but not so for low iron-high zinc aqueous solutions.
5.2.2.2. Large Scale Test
Intermediate scale continuous testing in the Bell system showed that
calcium was extracted concurrently with the zinc and that it precipitated in
the strip cells as gypsum. Procedural techniques were worked out to eliminate
the transfer of solid gypsum back to the extraction stages. Testwork was
conducted using a variable number of stages of extraction and stripping.
Phase I Study
Large scale testing was conducted in seven cells of the Reister testrack.
The procedure used in the testwork was developed on the small continuous Bell
system. The procedure consisted of: connecting the stages so that four
extraction stages and three strip stages were used (Figure 5.6); the cells were
loaded with the proper organic to aqueous ratio (0/A » 1 or 0/A • 3); solution
flow was initiated at the desired flowrate (up to 500 cc/mln.); samples were
taken (and pH monitored) of raffinate from stage two and stage four and from
the strip acid into and out of the system as a function of time;-"phase
Interfaces were observed for muck or crud formation.
Phase II Study
Large scale testing was conducted in ten cells of the Reister testrack to
Investigate a potential flowsheet allowing for low iron bearing solutions to be
treated by solvent extraction without prior jarosite precipitation. The
concept for the study was that iron could be removed from the leach solution at
low pH by D-EHPA (see Figures 8.10a, b); iron would then be stripped from the
loaded organic by a HC1 solution; the organic phase would then contact the
leach solJtion (at a higher pH) to extract zinc; the zinc loaded organic would
then be stripped by a sulfuric acid solution.
The procedure consisted of connecting the Reister testrack to provide a
flow pattern as presented in Figure 5.7.: one stage of iron loading; one stage
•
43
-------
SETTLER
MIXER
Aqueous
MIXER
SETTLER
•o
SETTLER
MIXER
«*-
.,-., 1
Iqntout
-*2aj2ie_
—
MIXER
SETTLER
?
0 J
Aqueous Feed
_ Aqueous
pll Adjust'
Aqueous
Raffinate
Pregnant Strip Crystal Hzatlon Return Strip Acid
Figure 5.6. Zinc solvent extraction flow pattern.
-------
Aquoout f«oo
Figure 6.7. Zinc and Iron solvent extraction.
-------
of sulfuric add stripping after iron loading for zinc removal (ferric ions are
not stripped by a 200 gpl H.SO. solution); two stages of iron stripping; three
stages of zinc loading; and three stages of zinc stripping. The cells were
loaded with the proper organic to aqueous ratio; the solution flowrates and
cell interfaces were established; and samples were taken of raffinate after
each stage of contact as a function of time. Phase interferences were observed
for muck or crud formation.
5.2.2.3. Organic Degradation Testwork
Long term load strip testwork was conducted in a Bell engineering 600 cc,
ten-stage continuous testrack. The testrack cells were connected to provide
one stage (low pH) extraction of iron; three stages (higher pK) of zinc and
iron extraction; one stage of sulfuric strip for zinc removal from the iron
loaded (small amount of zinc also loaded) organic; two stages of sulfuric strip
for the zinc loaded organic; and three stages of hydrochloric acid strip for
iron loaded organic.
The purpose of the testwork was to expose tne organic extractant to a long
term, many cycle load-strip sequence to determine whether the extractant was
degraded with use.
Potential degradation of the organic extractant was followed by closely
monitoring the element concentration in the raffinate solution and by
collecting organic samples after approximately every twenty liters of aqueous
contact. The organic sample was stripped twice wi'.h 200 gpl sulfuric acid then
exposed to a standard leach solution (two contacts). The effectiveness of the
organic phase extractant was determined by its ability to remove zinc and iron
selectively from the standard solution.
5.3. CHROMIUM OXIDATION
The oxidation of Cr*
chlorine gas (and other oxidizing agents) and by electrochemical oxidation.
The oxidation of Cr in the aqeuous phase was studied by exposure to
46
-------
5.3.1. Chromium Oxidation by Chlorine
5.3.1.1. Phase I Study
The oxidation of chromium by chlorine gas was studied first on a 100-500
cc scale; followed by 1.5-15 liter scale tests; then large scale tests ft
thirty liters and seventy-five liters. The procedure was to prepare a solution
with appropriate chromium content (usually prepared by kettle leaching a
sludge, removing the Iron by jaroslte precipitation, removing the copper by
LIX-622 solvent extraction, removing the zinc by 02EHPA solvent extraction);
adjust pH; purge in chlorine to establish a desirable solution Eh; sample as a
function of time to determine the extent of Cr+3 to C+6 oxidation.
Large scale testwork was performed In a 40-liter polypropylene vessel and
In a 120-liter polypropylene vessel. (A schematic representation of the
reaction system 1s presented In Figure 5.8). A chlorine lance was constructed
from PVC and the sparge rate adjusted to maintain the solution Eh at approxi-
mately 1000 mv. The experimental results are presented in Section 8.9.1.1.
5.3.1.2. Phase II Study
Large scale testwork was continued during the Phase II study using an
efficient chlorine gas-solution contactor system; a chlorinator. A chlorinator
1n Its simplest design resembles an aspirator system. Liquid solution Is
pumped through a venturi. Pressure change 1s generated that aspirates chlorine
through a side port. Turbulence Is created In the solution by movement through
chlorine gas. A schematic drawing of the chlorinator system 1s presented In
Figure 5.9. The experimental results and discussion of results are presented
in Section 6.36 and Appendix 8.9.1.1.2.
5.3.2. Electrochemical Oxidation
5.3.2.1. Phase I Study
Solution oxidation of chromium in an electrochemical reactor is depicted
schematically in Figure 5.10. Only small scale oxidation studies were
47
-------
Chlorine (Us Nitric AgiUtor. 1/4 H.f.
Steel
Liner
Figure 5.8. Chlorine oxidation reaction vessel (Cross section view).
48
-------
Agitator
Tank I
Figure 5.9. Schematic drawing of the chlorinator system.
-------
or
O
Cathode
Keobrane
Anode Cathode
SIDE VIEW
TOP VIEW
Anode Bus Bar
So
Figure 5.10. Electrocheoical oxidation cell.
ft
END V1EM
Frame
Membrane
•&»
Kerebrone Detail
-------
conducted In the Phase I study. The results were very encouraging and a large
scale reactor was Included In the second phase study. Two small scale studies
were performed; a series of batch oxidations and a continuous flow oxidation
study. A summary of the experimental results Is presented In Section 3.9.1.2.
The batch tests were performed on zinc raffmate prepared during the
sequential series five teslwork. The solution was. therefore, relatively free
of Fe. Cu. and Zn. The solution contained a mixture of chromium and nickel.
This solution was used to Mil the anode chamber (approximately one liter) and
a 180 gpl H2S04 So1ut1on was used to f111 tne cathode chamoers. The desired
cell voltage and current density were established and the oxidation allowed to
proceed for a designated time. Samples were taken as a function of time and
analyzed for all metal values and for Cr /Cr content.
The continuous test was conducted on the same zinc rafflnate solution.
The anode chamber was filled with zinc rafflnate partially oxidized previously
1n the batch tests ( 69% oxidized chromium). The two cathode chambers were
filled with 180 gpl H2S04' Unoxid*zed Zlnc rafflnate was fed continuously Into
the anode chamber at 3-5 cc/.n1n. and a similar volume was withdrawn. The exit
stream was sampled as a function of time.
5.3.2.2. Phase II Study
An electrolytic cell was constructed of 3/8 in. acrylic sheet material.
The cell dimensions were: 18 in. length, 12 in. width. 12 In. depth. Overflow
weirs were provided along the two sides of the cell. The overflow solution was
collected at the ends of the cell and was recirculated to the bottom of the
cell. The base of the cell was fit with a false bottom 1n the shape of an
Inverted acrylic pan one-inch high. Solution distribution holes (1/32 1n.
diameter) were placed on all four sides of the pan at 1/4
-------
by 10 1n. Nafion 423 diapnragm. The diaphragms were secured in place by
plastic flanges. A cross pipe was drilled with small holes and placed along
the length of each Nafion diaphragm. This arrangement allowed air to be blown
upward across the face of the diaphragm. The two chambers are depicted
schematically in Figures 5.11 and 5.12.
The electrochemical oxidation cell was formed by setting the smaller
chamber inside the larger chamber on the false bottom. The inner chamber is
the anolyte cell where oxidation occurs. The outer chamber is the catholyte
cell where reduction occurs. Anolyte solution containing the Cr and N1*Z
Ions is prevented frou intermixing with the catholyte solution containing
sulfuric acid by the walls and the Nafion diaphragms. Busbars for current flow
were made from 3/4 in. copper tubing. Current was supplied to the busbars by a
100 amp DC power supply (Lambda Model LES-F).
Electrode material was lead. Electrodes were 1/4 in. sheets by 12 in. by
4 In. (both cathodes and anodes). Later In the study special high surface area
anodes were constructed and investigated. These anodes were constructed using
a 1/8 in. perforated lead sheet. A 12 in. by 6 in. section of the perforated
lead sheet was laid down and layered with plumber's lead wool then overlapped
with a section of perforated lead sheet. The sides were folded over and
crimped to form a structurally strong anode.
Initial static tests (anolyte was not continuously fed into the anode
chamber) were performed using the lead sheet electrodes with an
anode:diaphragm:cathode ratio of 1:1:1. The applied voltcge was 3.5 v. The
Initial current densitj
over a 24 hour period.
9 2
Initial current density (c.d.) was 8 amp/ft . The c.d. Increased to 12 amp/ft
Anolyte solutions w»re sampled and analyzed fcr total chromium, hexavalent
chromium, and nickel. The hexavalent (oxidized form of chromium} was
determined by exposing an aliquot of the anolyte to an equal volume of Rohm and
Haas anionic exchange resin IR-900. The resin-anolyte mixture was shaken for
five minutes, then the solution was recovered and analyzed for chromium content
52
-------
Electro* Kiel
Ul
T
• IOC VIBW
— It»/« —
Inner Cell Petition
r
i
i
i
i
i
L.
I
• Cttnolrte Recycle Inlet
END VIM
16 1/2
Figure 5.11. Electrochemical oxidation
cell outer chanter.
TOP vin
-------
Nation
Air Sparge
Air Sparge
01ichor go
"N
• IK VIEW
IA \tn "
'v^
s
,
\
0
I
% 71/4"
1
|
Catholyt* RocycU
Oltporilon Ploto
VII
TOP VIEW
Figure 5.12. Electrochemical oxidation cell Inner chanter.
-------
resin). The difference between the total chromium in the original sample and
the chromium analyzed in the ion exchange resin treated solution was taken as
the hexavalent chromium content. A standard solution of cnromic acid
containing 20 gpl Cr was prepared. An aliquot of this solution was treated
similar to a test solution. Hexavalent chromium extraction by the resin in
five replicate samples showed 98.6 ^ 1.5 percent removal by the ion exchange
resin.
Catholyte solutions were analyzed for total chromium and nickel. All
solution analyses were performed using a Perkin-Elmer 303 atomic absorption
spectrophotometer using a nitrous oxide-acetylene flame.
The results and discussion are presented in Section 6.3.6. and Appendix
8.9.1.2.
5.4. CHROMIUM PRECIPITATION
The oxidation of chromium (Section 5.3.) resulted in a leach solution
-2 -1 +2
containing only chromium, as Cr.Oy or HCr04 . and Ni . The oxidized
chromium can be separated from the nickel cations by precipitation as lead
chromate. The lead chromate precipitation is a way of removing the chromium
selectively from the nickel and it also provides a means of concentrating the
chromium, i.e., the separated lead chromate-solid phase can be releached to
form a high concentration chromic acid and solid lead sulfate. The lead
sulfate then can be recycled to the oxidized leacn solution to precipitate more
chromi urn.
The experimental procedure used to consider the precipitation of chromium
consisted of small beaker tests to observe the effect of pH, time, and amount
of PbSO. on the recovery of chromium from solution. These tests were followed
by large scale precipitation experiments in an agitated vessel. The large
scale test procedure consisted of feeding a predetermined amount of lead
sulfate Into 45 liters of a leach solution previously sequentially treated for
Fe, Cu, and Zn removal; agitating the solution to suspend the
55
-------
tlon of time so that the degree of chromium removal could be determined; main-
taining the pH in the range 3.5-4.5; terminating the agitation to allow the
solids to settle (15-30 minutes); decanting most of the solution from the
solids; and recovering the PbS04-PbCr04 solids by filtration; redissolution of
the lead chromate in the solids in a sulfuric acid solution to determine the
ability to concentrate the chromium and to observe the contamination of other
metal ions in the resulting chromic acid. The results and discussion of re-
sults are presented in Section 6.3.7 and Appendix 8.10.1.
5.5. NICKEL RECOVERY
Nickel 1s the last metal 1on to be removed from solution. Its concentra-
tion In solution 1s usually In the range of 2 to 6 grams per liter. Therefore,
It must be concentrated. Two major means of concentration were investigated.
I.e., precipitation as nickel sulfide and solvent extraction.
•
5.5.1. Sulfide Precipitation
Sulfide precipitation was Investigated by small scale testwork utilizing a
design matrix to establish the Important experimental variables. Tests were
conducted in small beakers to establish the influence of pH, time, and NagS
concentration. The small scale testwork was followed by a large batch test on
42 liters of leach solution (pretreated for Fe, Cu, Zn, and Cr removal).
The large scale test procedure consisted of: feeding a solution of Na2$
slowly Into the reaction vessel; maintaining the solution pH in the range 4-4.5;
sampling as a function of time; agitating the slurry to keep the precipitated
nickel sulfide suspended In the solution phase; terminating uhe agitation and
filtering th-. solids from the solution. The results and discussion of results
are presented in Section 6.3.8 and Appendix 8.11.1.
5.2.2. Solvent Extraction of Nickel
Solvent extraction of nickel is not commercially practiced (except in
ammonlacal solutions). Therefore, only preliminary small scale tests were
56
-------
Conducted to Investigate potential solvent extraction concentration of nickel.
All of the testwork was performed In small (125-250 cc) separatory funnels.
The procedure used was the same as described for copper extraction in Section
5.2.1.1. The results and discussion of results are presented in Section
8.11.2.
57
-------
SECTION 6
RESULTS AND DISCUSSION
6.1. LARGE SCALE SEQUENTIAL TEST .MASS BALANCE (HIGH IRON)
A flowsheet summarizing large scale sequential experimental studies is
presented in Figure 6..1. Included are mass balances for Cu, Fe, Zn, Cr, Ni,
Cd, Al, and Ca. A summary of the distribution of each element into the various
products is presented in Tables 6.1 and 6.2. The metal content of each solid
product is presented in Table 6.3. The element distributions presented in
Figure 6.1 and Tables 6.1 through 6.3 are based on calculated values for 100
pounds of sludge and are, therefore, hypothetical numbers. Tne distributions
are, however, based on data generated in the large scale sequential testwork
presented in Section 8.13.
The throw-away product in the process is the leach residue-jarosite solid
mixture; i.e.. there are about 15,000 grams (33 pounds) of solids in the
starting 45,400 grams (100 pounds) of sludge; from the leach of this solid
material 4,800 grams of leach residue remain and 6,800 grams of jarosite are
produced. A large fraction of the iron (>95X) is rejected to the solid. Some
metal values are also lost to the solids; i.e., 10% copper, 61 Zn, 18* Cr. and
61 Ni. The copper loss is higher in the large scale testwork than noted in the
small scale testwork; nickel and zinc are similar to other testwork; and
chromium loss is quite variable'but usually falls within the range of about 15
to 25 percent.
•
The reason for the apparently high copper and chromium loss during the
jarosite precipitation process is related to the presence of phosphorus (note
the sludges in the Phase I study contained 2-41 phosphorus. Table 4.1). The
jarosite conditions are ideal for the partial deposition of copper and chromium
as phosphates; see Figure 6.2. The equilibrium chromium content (at 80°C) is
58
-------
en
vo
Figure 6.1. Treatment of
45.4 kg (100 Ibs.) of high Iron metal hydroxide sludge per day:
element distribution.
••••••••••••••••••••••••••••••••••••••••••UN
Voluie or
45.4 kg (100
•••••••••••••••t>*t ••••••••••••••*i «•••••••••• in tin
Nass Concentration
Fa Cu Zn Cr
1. Sludge (33t solids) : 15.0 kg solids kg 2.56 0.87 1.24 0.91
30.4 kg solution t 17.1 5.8 8.3 6.1
2. Recycle Solids (35* solius): 0.34 kg solids kg 0.13 0.00
-------
Figure 6.1. Continued
Volucc orlUst
6. Residue Solids : 4.8 kg (dry basil) kg
(not separated, i.e., sub- X
sequent jarosite precipi-
tation perfoned in pre-
sence of leach residue)
Concentration (kg/day or
f«
fmmit^^
0.21
4.3
Cu
•^•V^PV
0.06
1.2
Zn
^^•MV
0.06
1.3
Cr
^^i^m^tm
0.03
0.7
_NJ_
0.01
0.3
Cd
^^^^•v
0.00
0.0
ftl
^^•^••i
0.01
O.J
c*
••MW
0.25
5.3
7. KOH Solution (500 gpl)
8. HO (30X)
: 10.0 I
: 2.S I
S
Jarosite Precipitation
•85-92°C
•6 hrs.
•pH . 2.0-2.5
9. Leach residue - jarosite
Solid (6St solid): 17.8 kg
(11.6 kg solids.
6.2 kg solution)
Evaporative Solution loss: 24 I
**
fe Cu Jn_ _Cr_ Hi Cd »1 _Ca_
Eitractlons (t): 97.0 2.7 2.0 15.0 3.6 6.2 14.5 0.0
[IOIE: Chroiiui loss would be auch less if the jarositel
solids are releached with H2SO^ at a pll of< O.SJ
13 I wash water
Cu Zn Cr Hi
Cd
Ca
kg 2.50 0.08 0.08 0.16 0.02 0.00 0.19 0./5.
* 21.6 0.7 0.7 1.4 0.2 0.0 1.6 2.2
-------
Figure 6.1. Continued
• P I •••>•»••••• •••••••l»l«l»«l*»*»t«l»«IH ••••*••••••*•• >•
10. Filtrate
: 224 1
gpl
w
Solvent Extraction of Copper
•Initial pH - 1.7
•leip. - 40-50°C
'Iwo-stage extraction, O/A-1
•luo-stage strip, O/A-1,
180 gpl H^SO^
•15 v/o IU-622. 85 v/o
KIRNAC 4708
•250 cc/«in. each phase
pll .
••••••••••••P.*
12. Raffinate
13. KaOH («00 gpl): 1 liter
: 22<. I
gpl
Concentration (Kg/day or gpl)
f«
0.31
0.07
Cu
3.51
0.80
2n
5.09
1.16
Cr
3.10
0.75
jsu
1.36
0.31
Cd
0.00
0.00
Al
0.98
0.22
_Ca_
0.18
0.04
•*••••••••••••••••!•••!•••••••••«•••••••••••••••••••
Extraction Efficiency: Stage 1 - 96.8* Cu
Stage 2 (pH . I.5) - 95.3X Cu
II. Copper aay be electrouon (0.80 kg)
or
crystallite* as CuSO^SI^O (3.15 kg)
Concentration
fe
^••^P^MIV^HI
0.31
(90X f."*)
Cu 2n
0.005 5.09
Cr Mi Cd Al
3.30 1.36 0.30 0.98
0.07 ••••••••••••••••••••••I•••••••••••••••••••••••••••••••••••(•••(••(••••••••••••••••••••••••••I•»••••••••«•••«••••>••••
I
Ca
0.18
0.04
-------
Figure 6.1. Continued
o»
ro
Stage: 1 2
70. OX Zn 50.0X Zn
30.0X Al 20. OX Al
50. OX Ca 20. OX Ca
50. OX Fe 20. OX Fe
Stage: 3 *
70. OX Zn SO.OX Zn
30. OX Al 20. OX Al
SO.OX Ca 20. OX Ca
SO.OX Fe 20. OX Fe
Strip Efficiency
each stage : 85. OX Zn
S3. OX Al
100. OX Ca
• O.OX Fe
f>lllllllllll>l»l»ll»llllll«< «•!«•• •!•••••
Fe Cu
IS. Raffinate: 224 1 gpl 0.07 0.00
kg 0.02 0.00
Solvent dtractkon of Zinc
•Initial pH . 2.5
•leap. - 40-50°
•Four stages of eitracllon.
pH adjusted bach to 2.5
after second stage. O/A-1
•Ihree stages of strip
(200 gpl H2SOt), O/A-1 ••••••••••(••••••••••••••••••••••••••••••••••••••••••••••mi
•40 v/o Of IIP*. 60 v/o
KIRHAC 470B
••••••••••III!
14. Zinc lay be crystal! tied as ZnSO%-7H2b
U m i . Composition of solution from which Zn is
crystalliied:
Zn Al Ca Fe Cd
kg 1.14 0.15 0.03 0.00 0.00
(ppt. as gypsum)
iiii«*i«iiiiiiiiiiftii»ii«i«i««iii»iiiiiiii
-------
Figure 6.1. Continued
Oi
CJ
Chroaiua 0< Ida! ion
•Initial pH . 4-5
•leap. - 30-5u°C
•Retention liae •
4-5 hours
«
•
Oildatlon Efficiency: 85*
•
""M6. Precipitate (35X solids): .34 kg solid*
0.65 kg solution
fe Cu In Cr Hi CJ M
DH . 4 5 kg 0.02 — — 0.11 -- — 0.03
X 5.8 .. .- 32.3 .. .. 8.8
' fe Cu In Cr Hi Cd Al Cd
17. Oiidiied Solution: 228 1 gpl <0.l.
-------
Figure 6.1. Continued
o»
*
*
Chroiiu* Precipitation
•leap.: tabient
•li«: 0.5 hr.
•2X Stoichioiftric
Requirement of PbSO^
•Initial pll . 3.5-4
i Huh ?K fntr:
10 Dkr.n. _Dk«n. f7»»
.
pH - 3.5-*
fe Cu In
kg 0.00
t 0.0
laOH
ipped Solution: S.2 1
in ii ••••••i •••••••••••»• •• ••••••••••
tolids): 6.0 kg PbSO^-PbCrO^
7.6 kg lolution
Cr Hi Cd «l
0.6* - -- 0.03
11.0 00 00 0.5
••• • ••••«•
Ca
--
•••••••••••••••••••••••••••••••••••••••••••••••••••••••••••••••••••••••••••••i
QftVVB p B B •) •) V p1 p1 V B)pl)t)AllflllflllftVAWplpfl0p'tttfflftp! IB kfppwppwv p • • p p w v
70. Solution: 730 1
WHVVFV'VUPVVVVVBWtVPVBTHVVVBHVWVH
fe Cu in
gpl <0.l.
-------
Flqure 6.1. Continued
o\
Ul
Bickel Precipitation
Mt.p. . 2S-3S°C
•list - O.S hr
•21 Stolchloietrlc
Requirement of lltjS
(not optitiiod)
•Initial pH . 3-4
PH . 3-4
23. Recycle Solution (to leach and lo
Hater cakeup): 231 I
lo flecycle
• O.S I 400 gpl
«•••••••••••••••••••»•••••«•••«•••••• *••••• Miff •••••!••••••• •••••••!••• !•••
-» 22. Sulfide Precipitate (3St tolidt): O.S kg tolidt
1.4 kg solution
111 Cr 2n 01
kg 0.32 <0.00 0.02 <0.00
* 64.0 — 4.0
Concentration
gpl
-------
TABLE 6.1. TREATMENT OF METAL HYDROXIDE SlUDGE: ELEMENT DISTRIBUTION SUMMARY
Oi.tr Ib.litn (k«/di»)
Input
Cut
•r Product Rill (kg)
llgart I.I)
Volun (1)
ft
Cu
tn
Cr
II
Cd
II
Co
Stroit •».
1.
2.
I.
4.
S.
1.
1.
1.
9.
10.
II.
17.
11.
It.
II.
II.
II.
Id.
19.
70.
71.
77.
71.
Slvdgt li.O-.olldt
IfCfClt Solidi O.H-iolid.
•tcyclt Solalto*
•2S04 Icid
Itick Solution
•t.idut Solid* 4.1 kg
•OH (400 gpl)
NjOj (MX)
Ittidut Solid*- ll.t-tolldt
Jirotltt
mtrott
Copptr Strip
Clrt.lt
Cn RifHiulo
•lOH (400 «pl)
tine Strip
Circ.lt
line RoUimtt '
PrtclplUlt O.lt-tolldt
(tilt •* «7)
OildKtd SoUtio*
lud Sulfilt I.I kg
•bCrOi-PbSO^ l.0-*olld
Sold ion
••|S Sol.tlon
(»s wD
Svlhdt PrtclplUto 0.4-*ollJ
ticrclt Mml Sol*.
M.4 kg-*olulloB
O.llkg-iolutio*
IM.
10.
779.
10.
7.
1
71t.
1.
1.
2It.
1.
II.
II.
I
|
1
1
1
kg-tolotlon
1
1 <«to»i
1 Orginic
Ilt.O 1
0.65 kg-MlulIra
221.0 1
7.1 ka.-tol.it loa
2JO.O I
1.0 1
1.1 kg-iolallaii
21.1 1
7.U
O.I)
*o t
l.ll
0.71
I. M
0.01
0.01
0.04
0.07
0.07
0.00
0.00
0.00
0.00
<0.l.
O.I!
0.00
CO 1
0.17
O.Ot
0.01
0.10
0.10
-------
o»
TABLE 6.2. TREATMENT OF METAL HYDROXIDE SLUDGE: DISTRIBUTION TO SPECIFIC
Distribution To Specific
Product
Leach Residue-Jarosite
Copper SX Circuit
Zinc SX Circuit
Chromium Slurry Oxidation Solid
(Recycled to Leach)
Lead Chromate-Lead Sulfate
Sulfide Precipitate
PRODUCTS
Distribution Ci)
Fe
97.6
0.0
2.0
0.8
0.0
0.0
Cu
9.2
92.0
0.0
0.0
0.0
0.0
Zn
6.4
0.0
91.9
0.0
0.0
1.6
Cr
17.6
0.0
0.0
12.1
70.3
0.0
Ni
S.9
0.0
0.0
0.0
0.0
94.1
Al
42.2
0.0
33.3
7.1
6.7
0.0
Ca
86.6
0.0
10.0
C.O
0.0
0.0
Notes: . Distribution balance based on flowsheet Figure 6.1.
. Detailed experimental results for large scale sciential testwork presented in
Section 8.13.
-------
O»
CD
TABLE 6.3. TREATMENT OF METAL HYDROXIDE SLUDGE:
Product
Starting Sludge (Solids)
Leach Rest due
Jarostte
Lead Chronate-Lead Sulfate
(27.91 PbSO,. 68.31 PbCrO..
l.SX AI(OII73.
Nickel Sulflde
ELEMENTAL CONTENT IN SOLID PRODUCTS
Elemental Content (t)
Fe
17.1
4.4
33.7
0.0
0.0
Cu
5.8
1.2
0.3
0.0
0.0
Zn
8.3
1.2
0.3
0.0
0.4
Cr
6.1
0.6
1.9
11.0
0.0
HI
2.3
0.2
0.1
0.0
64.0
Al
2.8
0.2
2.6
O.S
0.0
Ca
2.0
5.4
3.8
0.0
O.U
Notes: . Based on flowsheet Figure 6.1. /
. Detailed experimental results for large scale sequential tes^ork presented In
Section 8.13.
isu
-------
Concentration,
log (ppm)
I I
I
COPPER, 25°C
CHROMIUM. 80°C
LCHROMIUM. 25UC
I I I
1.0
Solution pH
2.0
Figure 6.2. Solubility of chromium and copper phosphates.
69
-------
cbotrt 0.5 gpl at pH of 2.C. The effect of pnosphate precipitation is not
considered a deterent because tne jarosite (once formed it does not readily
redissolve in acid solutions) can be relcached to redissolve tne chromium
phosphate and copper phosphate.
If an operating plant has phosphorus containing sludges then an acid
releach (pH and temperature controlled) of the jarosite may be desirable. The
resulting leach stream could be fed into the solution stream from the jarosite
filtering unit operation. This approach is discussed in Section 6.3.2.
Other investigators have reported chromium contents in potassium
jarosite' ' ' to be in the range 0.6-1.61. The present results show 1.9% Cr.
It 1s presently not clear whether this loss is a true chromium substitution for
Iron to form K(Fe,Cr)3(OH)g(S04)2 or whether a coprecipitated chromium
phosphate phase forms on the jarosite surface. The condition of the sequential
tests was very oxidizing. This also has been noted in the present work to
enhance chromium loss. It is reported ir: the literature ' that CrO.° may
substitute completely for SO^* in the jarosite structure, i.e., a
KFe3(CrO.)2(OH)g compound forms. Therefore, under high'iy oxidizing conditions
the following reactions are expected to occur:
Chromium is oxidised slowly,
+3
2Cr
= 2HCrO"4 + 8H* c ° • 0.6 volts
Iron also oxidizes
H202 + 2Fe*2 + 2H* » 2Fe*3 + 2H20 e
1.0 volts
Both reactions are thermodynamically feasible. As long as there is any
ferrous icn present the HCrO." ions will oxidize the ferrous io.is:
HCr04" + 3Fe*2 + 7H*
3Fe*3 + Cr*3
e « C.4 volts
70
-------
When the iron has all been oxidized, HCrO." (present at all pH levels if
/ c\ *
chromium content is less than = 1 s?ll '), then should form. This oxidized
chromium is. therefore, available for reaction to form the jarosite.
Therefore, proper solution conditions must be chosen to minimize chromium loss,
I.e., the addition of a minimum amount of oxidizing agent is required
(sufficient to oxidize the iron but not the chromium). Also significantly less
loss of chromium can be expected in those systems that are relatively low in
Iron content, e.g., if 0.6-1.6% Cr* is incorporated in the jarosite
precipitate* ' then if the Iron content in a solution is 1 gpl (and the
chromium level remains at 3.8 gpl as illustrated in Figure 6.1.) Instead of ten
grams per liter the loss of chromium to the solid would drop to the range 0.1
to 0.41 or 'ess. This conclusion needs further testworfc for verification but
kettle test results support the conclusion. Further discussion of impurity
incorporation in precipitated jarosite is included in Appendix Section 8.3.1.
A list of summary comments for each large scale unit operation 1s
•
presented below. A detailed presentation and discussion cf all large and small
scale test work are presented in the following section, 6.3, and in Appendices
8.2-8.16.
"The sulfuric acid leach operation is effective in redissolving the
metal values. The dissolution is rapid and without control
problems. The leach is carried out in a single 270 liter vessel.
The conditions required are well characterized, and rather mild,
I.e., one-half hour, 40-50 C, sludge/liquid ratio of 0.8, acid
content to control pH in «.h' -ange 0.5-1.5, and agitation
sufficient to suspend the particulate in the solution phase.
The sludge dissolution is essentially complete in less than
one-half hour. Therefore, the leach operation is not the
controlling step in the overall treatment sequence. The leach unit
operation is capaole of treating over a ton of sludge per eight
hour day. The filterability of the leach rzsidue product is
difficult. The filterability of a mixed leacn residue-jarosite
product is rapid and effective. Therefore, in most testwork the
jarosite precipitation process was performed in-situ with the leach
residue solids.
'The iron removal unit operation is via the precipitation of
potassium jarosite. The precipitation process requires elevated
temperatures and relatively long reaction times. Two hundred
liters of leach solution slurry can be treated in six-eight hours.
71
-------
The jarosite process allows iron to be removed from an acid
solution. The product is a crystalline compound tnat has excellent
settling and filtering properties. Tne {ron removal process has
been demonstrated on high iron sludge materials, i.e., 15-?0» iron
in the starting sludge solids. This means tnat for these
particular sludges a significant quantity of leach residue-jarosite
solids are formed, e.g., 11.6 kg of solids or 17.8 kg of wet .
material (see Figure 6.1.) for a 17.IS Fe bearing sludge iraterial.
Therefore, the disposal of 17.8 kg would be required instead of
45.4 kg or approximately forty percent of the original sludge
weioht. A significant quantity of sludge material exists that has
iron contents much lower than the above values. The jarosite
process is also effective for treating the low iron containing
sludges, e.g., two-four percent iron. The quantity of leach
resid^e-jarosite solids produced from such sludge material would be
rather small, e.g., a sludge similar in composition to the Figure
6.1 material but containing two percent iron would yield 5.6 kg of
leach residue-jarosite solid. This quantity of solids translates,
at 65% solids, to 8.6 kg of disposable material instead of 45.4- kg
or approximately one-fifth the original sludge weight.
Jarosites are widely produced in the zinc industry. They are
deposited in lined storage ponds. It is difficult to state whether
their heavy metal content means that the jarosite should be
considered a hazardous material but even if t^at is the case at
least a significantly smaller weight of material must be considerr.-d
for disposal.
High iron sludges do (low iron sludges do not) present a problem
for chromium recovery. Significant aiiounts of chromium are lost
when the jarosite precipitation is performed. It is believed that
the loss can be minimized by maintaining conditions such that
chromium is not oxidized and the &H is maintained below 2.5. A
releach of the jarosite solids appears to be necessary, if the
sludge is a phosphorus containing sludge, to prevent both chromium
and copper loss.
Mechanical control of tne system is no problem. Chemical control
must be exercised to ensure that the pH is maintained in the range
1.8-2.5 and that the iron is in the ferric form. Solid-liquid
separation is effectively accomplished by allowing the leach
residue-jarosite to settle; decanting the solution from the solids;
and pumping the small volume of renaming slurry to a filter press.
•The removal of copper is accomplished by solvent extraction (SX).
The extraction of copper from zinc, chromium, nickel and aluminum
1s selective and effective (>96% extraction per contact stage).
Copper contents of a few mg/liter are achievable in two stages of
contact and one stage of strip. The pK of the solution exiting the
jarosite precipitation unit operation can be treated without
adjustment.
72
-------
The SX testrack is designed to treat up to 200 liters cf solution
per day. The design throughput is 500 cc/min. for each phase.
Most tests to date have been performed at 250 cc/min. Ten contact
mixer-settler units a>*e available for copper SX. Therefore, this
unit operation is not the slow step in treatment sequence, i.e.,
three concurrent streams could be treated (each in three cells) at
one time; at 250 cc/min., 363 liters could be treated per day. .
Large scale testwork has been conducted for up to six hours. Control
of flowrate and interface levels is easily achieved and requires
constant attention only during initial loading of the system. Once
the system interfaces have been established little operator atten-
tion is required.
"The removal of zinc is accomplished by solvent extraction. The
extraction of zinc from chromium and nickel is selective. Ferric
Iron, aluminum and calcium are partially coextracted with the zinc.
The extraction of iron (only between 0.2-0.6 gpl present) with zinc
is desirable because it provides a way of removing residual iron
from the solution. The iron once extracted into the organic is not
stripped by H2S04 but is stripped by HC1 acid (4-6N). Zinc is
stripped by ^$04 (200 gpl). Therefore, a means of bleeding Iron
from the process stream Is to load iron and zinc into the organic
phase, strip the zinc by contacting with ^04 (200 gpl) followed
by stripping the iron from the organic by contacting with KC1 (4N).
Both strip solutions can be recycled until the metal content is
appropriate for recovery of zinc as zinc sulfate heptahydrate and
for disposal of iron as ferric chloride solution.
Calcium 1s coextracted with zinc but poses no problem because it
precipitates as gypsum in the ^$04 strip circuit. It can be
effectively filtered continuously from the solution during solu-
tion crystallization of zinc sulfate.
The zinc SX testnck is the same design as used for copper removal.
Ten SX cells are available. The removal of 5 gpl zinc requires
four stages of extraction, three stages of zinc stripping, and one
stage of iron stripping. Therefore, the removal of zinc is the
limiting step in the present treatment process. Two hundred liters
can be treated at a flow rate of 400 cc/min. for each phase in an
eight hour period. Some flexibility does, however, exist by control
of the extracting reagent concentration in the organic and by changing
the organic to aqueous ratio in the system.
Control of flowrate and Interface levels *s easily achieved and does
not require constant attention once tne initial loading and interface
levels have been established; i.e., operator attention is
73
-------
minimal. The system can be shut off and restarted without
difficulty. Chenical control of pH'is required in zinc extraction
to achieve effective zinc remo.-al. Solution pH control is
exercised by adjusting pri after the first two-stages of contact.
Temperatures in the range of 40-55 C are desirable for rapid phase
disc.'oanement.
'Chromium removal is accomplished by first oxidizing chromium with
chlorine gas, then precipitating lead chromate. The oxidation is
more rapid if tne chromium is present as chromium hydroxide and the
system pH is maintained above four. Effective and rapid oxidation
on a small laboratory scale has been accomplished. The reactor
design used for large scale work was not as effective. Four to
five hours of contact were required in the large scale testwork.
Snail and large scale testwork has also been performed using
electrochemical oxidation in a partitioned electiode chamber-cell.
The results were encouraging and should be pursued further.
Chromium removal 1s very effectively achieved by precipitation of
the dichromate anions using lead sulfate. The removal of chromium
Is selective over nickel, i.e., nickel cations are not
coprecipitated with the chroinium. The prjcess is one in which lead
sulfate 1s regenerated for reuse, i.e., lead chromate can be
redlssolved In sulfuric acid to form chromic acid while
reprec'.pitating lead sulfate. The precipitation zf lead chromate
from tne oxidized leach solution is very rapid ( one half hour).
The lead chromate product is crystalline and dense. It settles
rapic'.ly and the solid-liquid separation is very easy and rapid.
Mechanically the system operates easily. Chemical control is
required to maintain the pH in the range 3.5-4.5.
'Nickel 1s removed by sulfide precipitation. The reaction is rapid
and near quantitative removal is possible. .The pH is maintained in
the range 4-5 to ensure that hydrogen oulfide is not released. The
solids are readily filterable.
In actual practice a deficiency of sodium sulfide would be used,
I.e., less than the stoichiometric requirement to completely
precipitate tne nickel. This procedure would leave some nickel in
the solution but the presence of nickel is not a disadvantage
• because the final solution is recycled to the leach unit operation.
Several alternate nickel recovery processes are possible.
6.2. LARGE SCALE SEQUENTIAL TEST MASS BALANCE (LOU IRON)
A flowsheet summarizing large scale experimental studies for low iron
bearing solutions is presented in Figure 6.3. Included are mass balances for
Cu, Fe, Zn, Cr, Ni, Cd, Al and Ca. The major difference in this flowsheet and.
74
-------
Figure 6.3 Treatment of 4.5 kg (100 Ibs.) of low iron metal hydroxide sludge per day: element distribution
45.4 kg (100 poundi)/day
••••••••••••••I •••••••••••••••••••••••••••••••••••••i
Voluie or Nass
!••••••»•••»••••• •••••••••*•••••••••• •••»•• •!•••••!••
Concentration (kg/day or
i ••••••••••••«••••••••••••• •• mill
X)
fe Cu Zn Cr Hi Al Ci
1. Sludge (25.0 solids) : 11.4 kg solids
34.0 kg solution
2. Recycle Solids (3SX solids): 0.5 kg solids
0.9 kg solution
3. Recycle Solution : 188.0 1
4. H2SO^ Acid : 6.1 1 (25.) Ibs.
Acid I
•40-60
•O.S h
•pH . 1
J
••••••••••••••fll«lll««l»**ll«ftl« •••••••••• ••••••41*
Voluie or Nass
kg 0.66 0.68 0.91 I. 52 2.
X S.b 6.0 8.0 13.3 17.
kg 0.01 0.00 0.0 0.21 0.
X 1.6 0.0 0.0 35. 0 0.
kg 0.0 0.0 0.0 <0.01 <0.
I
1 »«••••••«•»*• •••*••«•• •••••«•••••• •« >•••••• »»»I«*Il !••>••• •»•• I ••••••••••••••••••••••••••••••••t«l»l«l
1 96.5 9b.« 96.9 15.0
VICIIIIIIIIIMIIIIIIIIIIIIII •••••••
Concentration (kg/day or gpl)
fe Cu 2n Cr Ni Al Ca
5. Loach Solution : 248 I
6. Residue Solids : 3.2 kg (dry basis)
gpl 2.44 2.57 3.48 5.87 7.
0.61 0.63 0.86 1.46 1.
kg O.OS 0.04 0.04 0.06 0.
X 1.7 1.3 ' 1.4 1.8 2.
65 1.24 0.05
95 0.31 0.01
08 0.01 0.08
6 0.3 2.4
j ' ".
-------
Figure 6
(70* solid*) 3.2* kg solid «_ IU"P .^
1.39 kg solution ,"* .xf: . .
, ... " . . Solids .X^ liquid
(composition as above: Stnaa 6 ^s'
pH • 1.5
.3. (Continued)
11 1 Mash Mater
Concentration (kg/day or gpl)
ft Cu Zn Cr Hi
7. filtrate : 258 1 gpl 2.33 2.46 3.33 5.61 7.50
kg 0.60 0.64 0.86 1.45 1.94
• •••••I II I III III! •• ••••• *••••••••• ••••••••••••••••*••• !•• I !•••••••• Ill ••••• I II II •••!•• •••••••• ••••••••••• 1 •••••••
| 8. 0.6 NaOH (500 gpl)
^i Solvent Citraction of Copp
°* -Initial pH - 1.75
•Teap. ' tO-50°C
•three-stage entraction,
0/A . 1
*1vo-*tage strip, 0/A • 1,
180 gpl MjSOi
•15 v/o UX-622. 85 v/o
KCRHAC 470R
•250 cc/ain. each phase
* pH - 1.3
• »•••••••••• •••••••••• •!•••••••• ••••••••••••• •••••••••••ttllllllll
tr
••—Extraction Uficiency: Stage 1 (pll - I
Stage 2 (pH • 1
Stage : (pH - 1
Overall
• ••••••• •• •••••• in •••••••••••••••••••••••
» 9. Copper aay be electroiion (0.73
or
ciystalliied as CuSO -5H 0 (2.
Concentration
fe Cu Zn Cr Ni
10. Raffinate i 260 1 gpl 2.33 O.Ot 3.33 5.61 7.50
kg 0.60 0.01 O.B6 1.45 1.94
Al Ca
1.19 0.05
0.31 0.01
• •••••••••••••••••••••••••in
.75): 92.1k
.50): 85. 2S
.30): 43. 8%
: 99.3*
1-9)
13 kg)
Al Ca
1.19 0.05
0.31 0.01 '
-------
Figure 6.3. (Continued)
•••••••I•••
II.
i 11. 3.0 I laOH
Solvent Eitraction
of tine and Iron
•teip. - 40-50°C
•four-it «ge dtractlon 0/A • I
•Ihret-ttage HjSO^
•40 v/o OEHPA. 60
Strip
v/b KCRMC 510
•250 cc/ein [ach Phase
•HC1 (6 N)
•HjSOi (?00 gpl)
•pit 1.1 Into Cell One
•pll Adjusted to 2
••••••••••••••••••••••••••••••••••••••••••
n«t« : 26) 1
•• •*• •••••••••••••••••••••••••••••••••••••
Into Cell Inn
pH - 1.3
••••••••••••••••i
fe
gpl 0.01
kg 0.01
(500 gpl)
Citractien efficiency pH
fa In Ca Al
Coll 1 80.0 15.0 13.8 21.0 1.1
2 82.0 88.0 10.9 59.0 2.0
3 55.0 65.0 46.7 28.0 1.5
4 25.0 10.0 41.5 18.4 1.1
Overall 98.7 98.2 80.0 81.5
"* 12. Iron at led] In HCI (tie Section 8.4 ): 0.59 kg fe
Zinc at InSO^ in l^SO*:
Solution Composition:
2n Ai Ca
kg 0.84 0.25 0.01
••••••••••!••••••••••• »•••••••••••••••••••••••••••••••••••••••••••••
•••d •••••&••••••••• •••'§ •••••••••••••••!••••• !•••••• ••••••••••••••• fllVP**
Concentration
Cu In Cr Hi _*!_ Ca
0.02 0.06 5.60 7.50 0.22 0.00
0.00 0.02 1.45 1.94 0.06 0.00
• •••••• •!••••• •••••••••••••*•••••• •••••ajptJiipg ••••••«•••••••• «••• »•>•••••)•••)•••••••••• ••••p
' It. 8 1 500 gpl laOH
Chrooiu* Oildation
•Initial pH - 4-5
•leap. • 30-50°C
•Retention liie • 4-5 hrt.
Oildalion efficiency 65*
• •ll*lllll»l*l«l«lll«IKIIIf ••••••Itll«ll**lll»ll>l»*«*«««««>lll»
pH . 4-5
15. Precipitate (15* solids): 0.5 kg solids. 0.9 kg solution
ft Cr
0.01
2.0
0.15
10.0
0.02
4.0
-------
Figure 6.3. (Continued)
1
16. Oiidiitd Solution : 270 1
fe
gpl 0.00
kg 0.00
Cu
0.00
0.00
In
0.06
0.02
Cr
*.79
1.10
•1
7.28
1.9*
M
O.I*
0.0*
Ci
0.00
0.00
17. lead SulMtfe (it SloitMo-
•etrie requirement) : IS.16 kg
•«••••••••»••••!•«•••!•••••••!••ti••••••
I pH . 1.5-*
20. Solution
Chroaiue Precipitation
•fe«p.: Aabient «— 18. 1 liter 500 gpl laOH
•Hat: O.S hr.
•21 StoichioitlrU
Requirement of PbS04
•Initial pll - 3. 5-*
••••••••••••••I
: 260 liter*
! |9. PbCrO "-PbSA (70k $4lidt)i IS. 7 kg PbSO -PbCrO
pH . 3.5-*
Cr M •
kg 1.10 0.0*
« 8.1 0.1
Concentration
f» Cu lit Cr Ni Al Ca
gpl <0.l. Solution (12S gpl) : 8 I
-------
Figure 6.3. (Continued)
Nickel Precipitaticn
•leap. - 2S-3S°C
•ti«e . 0.5 hr.
•IX Stoichloaetric
Rcquireient of NajS
(not optiiiied)
•Initial pH - 3-*
pH • 3-*
23. Recycle Solution (to leach and to
H?ter laheup ):292 liters
0.5 I 400 gp)
• •• ••»••»••••••«••••••• III »• !••••••••• •••••••••••!•••••• .•••«••••••••>•>•>»•
22. SulfHe FrccipiUtf (35* tolldsj: 3.0 kg »oUdt
6.6 kg solution
•1
Zn
M
1.92 0.00 0.02 0.00
64.0 — 0.7
Concentration
Cu
<0.l.
-------
that presented in Figure 6.1 is that jarosile precipitation is not required.
Iron is removed from the system solution by solvent extraction using O.EHPA as
the extractant with subsequent ferric ion recovery from the organic phase by
stripping with hydrochloric acio solutions.
A summary of the distribution of each element is presented in Tables 6.4
and 6.5. The metal content of each solid product is presented in Table 6.6.
The distributions are based on data generated in the large scale and continuous
testwork presented in Section 8.4.
A major advantage of this flowsheet over the high iron flowsheet is the
elimination of the jarosite precipitation unit operation. Therefore, copper
and chromium loss does not occur and less disposable solids are created.
The throw -way product is the leach residue; i.e., there are about 11,400
grams (25.0 pounds) of solids in the starting 45.400 grams (100 grams of
sludge); from the leach of this solid material 3,ZOO grams of leach residue
remains for disposal.
A list of summary comments for each large scale unit operation is
presented below. A detailed presentation and discussion of all testwork are
presented in Section 6.3 and Appendices 8.4 and 8.8.
'The sulfuric acid leach operation for the low iron bearing sludge is
the same as presented previously, p. 71 .
'Solid/liquid separation of the leach residue can be successfully
accomplished by use of a filter aid, e.g., Udylite Oxyfin 985.
Pressure fiiteration is ineffective (the filter cloth plugs) in the
absence of a filtering aid.
'The removal of copper is accomplished by solvent extraction as
described previously, p. 72 .
"The removal of iron is accomplished by solvent extraction of iron
with D?EHPA in the first stage of the zinc extraction testrack. The
pH of the aquecus phase is decreased to approximately 1.0, then
contacted with a forty volume percent D-EHPA - 60 volume percent '
KERMAC 5103 kerosene. Iron is extracted leaving the zinc, nickel,
and chromium in the aqueous solution. Some zinc is coextracted but
80
-------
TABLE 6.4. TREATMENT OF METAL HYDROXIDE SLUDGE (LOW IRON) : ELEMENT DISTRIBUTION SUMMARY
Input
or Product
(tea figure 6.1)
Slrean lo.
1. Sludgt
. Recycle Solid*
. Recycle Solution
m
.
.
.
.
.
10.
II.
12.
11.
1*.
IS.
16.
17.
IB.
19.
20.
21.
22.
21.
H2SOt Acid
LeacK Solution
Retidu< Solid*
filtrate
RaOH (SOO gpl)
Copper Strip
Circuit
Cu RaFlinato
Ia3ll (500 gpl)
line and Iron
Strip Circuit (II
line and Iron
RafFinata
RaOH (SOO gpl)
(taoc ai 11)
Oddiicd Solution
lead Sulfate
I« 1.*
2S6.0
0.6
I.S
I.S
260.0
1.0
II. S
.1 lit. .rg) '•*
22*. 0
g-iolutlon
Aqucout
Organic
H?SOjJ01gpl)
HCI (61)
6.0 1
O.S tolldi 0.9 kg-tolullon
278.01
IS.2 kg
1.0 1
IS.? tolid 6.7 kg-iolulion
292.01
8.0 1
1.0 tolld S.6 kg-iolutlon
259.0 1
0.61
O.OS
0.60
0.60
0.00
O.S9
0.01
0.01
0.00
0.00
0.00
0.00
CO.L.
Cu
0.61
0.00
0.61
0.0*
0.61
0.61
0.01
0.00
-
0.00
0.00
0.00
0.00
0.00
0.00
-------
TABLE 6.5. TREATMENT OF METAL HYDROXIDE SLUDGE (LOU IRON):
Distribution to Specific
Product
Leach Residue
Copper SX Circuit
Zn and Iron SX Circuit
Chromium Slurry Oxidation Solid
(Recycled to Leach)
Lead Chrome te-Lead Sulfate
Sulflde Precipitate
DISTRIBUTION TO SPECIFIC PRODUCTS
Distribution (X)
Fe
7.6
0.0
90.9
1.5
0.0
0.0
Cu
• W^B^^^Vi^M
5.9
92.6
1.5
0.0
0.0
0.0
Zn
5.5
0.0
92.3
0.0
0.0
2.2
Cr
3.9
0.0
0.0
9.9
85.6
0.0
HI
1 Vl^iVBM^MB*
3.9
0.0
0.0
0.0
0.0
95.1
Al
3.1
0.0
78.1
6.3
12.5
0.0
Ca
88.9
0.0
11.1
0.0
0.0
0.0
NOTES: •Distribution balance based on flowsheet Figure 6.3.
•Detailed experimental results for large scale sequential testuork presented In
Section 8.13. T
-------
00
TA3LE 6.6. TREATMENT OF METAL HYDROXIDE SLUDGE (LOU
Product
Starting Sludge (Solids)
Leach Residue
Lead Chromate-Lead Sulfate
48.31 FoSO.. S0.4X PbCrO*.
1.3 A1(OH)3
Nickel Sulflde
IRON):
ELEMENTAL CONTENT IN SOLID PRODUCTS
Elemental Content (S)
_Fe^
5.8
1.6
0.0
0.0
Cii
6.0
1.2
0.0
0.0
^n^
8.0
1.6
0.0
0.7
Cr
13.3
1.9
8.3
0.0
Ni
17.9
2.5
0.0
64.6
Al
2.8
0.3
0.3
0.0
Ca
0 8
2.5
0.0
0.0
NOTES: -Based on flowsheet Figure 6.3.
•Detciled experimental results for large scale sequential teslfcork presented In
Section 8.13. f
-------
is stripped by 200 gpl sulfuric (iron does not strip). The iron
bearing organic pnase is then stripped with 6 N HC1 and returned to
the testrack for contact with tne aqueous phase (at pK 2-2.5) for
zinc loading.
The hydrochloric acid strip solution effectively removes the iron
from the O.EHPA but a problem witn this approacn is the relatively
large quantity of strip acid required. The hydrochloric acid
solution will only load iron to 5-9 gpl iron. Therefore, HC1
recovery and recycle would be necessary in a commercial operation.
Hydrochloric acid can be recovered from tne strip solution by an
additional solvent extraction unit operation. Recovery of hydro-
chloric acid was not investigated in this study but is practiced
commercially by Tecnicas Reunidas Company at its Espindesa plant.
The removal of iron from the leach solution is effective; solutions
containing <50 ppm iron can be produced. In the early stages of the
study of this flowsheet crud formation in the first contact mixer
and settler was a problem. In iron-phosphorus bearing solid phase
developed. The use of a low pH in the first contact ce". 1 and a low
aromatic kerosene eliminated this problem.
"The removal of zinc is accomplished by solvent extraction. The zinc
and iron solvent extraction system is one continuous system. A
large fraction of the iron is loaded in the first stage of the
testrack at a pH of about one. Zinc extracted into the organic
phase of cell one is stripped by contact with 200 gpl H_SO.; then
Iron is stripped from the organic by 6 N HC1. The organic stream
then then enters the second loading cell where it contacts the
aqueous leach solution (at pH => 2-2.5). Zinc and iron are loaded
into the organic in three stages of contact; then stripped in three
subsequent stages of sulfuric stripping.
Comments presented previously, p. 73 , apply to zinc solvent
extraction.
'The unit operations for the removal of chromium and nickel described
on page 73 to 74 are applicable also to this flowsheet reference.
6.3. UNIT OPERATION STUDIES
The discussion material presented in this section will be a summary of
results. Tables and figures will be presented to support each unit operation
summary. Support data and detailed discussion of experimental results and
discussion of alternate treatment possibilities are presented in Appendices
8.2-8.14. Some studies presented in the appendix section are not discussed
here. These were studies performed to guide the research team in their
84
-------
selection of the most appropriate flowsheet but wnose result: wera not
favorable enough tc warrant further consideration.
The experimental approach and philosophy for the laboratory verification
studies Include preliminary test of concept by screening experiments;
development of a two-level factorial design matrix for the experimental bench
scale studies; execution of the studies in the design matrix to establish which
variables are most important aid what the relative effect of each particular
variable is on the measured rasult; and subsequent use of the design matrix
effects (by using the Box-Hi Ifon "steepest ascent" approach) to optimize the
selection of experimental variables for further larger scale testwork.
6.3.1. Leach Studies (Detailed discussion and data presented in
Appendix 8.2)
6.3.1.1. Preliminary Testwork (Phase I)
Mixed metal sludge material was supplied by three sources in the Seattle,
WA. area; i.e., two electroplating firms and a chemical disposal firm. Sludge
compositions are, of course, variable and depend on many factors; such as
electroplating activity at a particular plant at a particular time; mixing of
spent liquor streams, etc. An illustration of composition variability between
sources and even within a particul\r source wac presented previously in Tables
4.1 and 4.3.
Three leach concepts have been considered, i.e., sulfurlc acid leaching,
cnlorine gas oxidative leaching, and caustic leachinr. Sulfuric acid leaching
will be discussed in this section; oxidative leaching and caustic leaching are
discussed in Appendix 8.13.
The sludge materials used in this stud} are designated by barrel number.
All materials used In the experimental program were mixed metal sludges
containing approximately twenty-thirty weight percent solids. The sample
preparation procedure used to prepare sludge for tescwork was: withdraw a 500
gram sample; mix and blend; sample at the time of a designated test to
determine moisture content; chemically characterize the starting sample;
85
-------
withdraw a specified weight of sample from the 500 gram batch for leach
testing. Experimental reproduceoility of the starting sludge solid composition
for a specific blended sample is presented in Table 6.10.
Sulfuric acid is a very effective leaching agent for treating mixed metal
sludge material. Based on design matrix and optimization studies a standard
Icuch was chosen for all subsequent leach tests, i.e.. the leach conditions
used in a majority of the subsequent testwork were: \ld. hour exposure;
agitation to completely suspend all particles in the solution phase;
temperature, 45-55°C; solid/liquid ratio. 200 gm sludge/250 cc added solution;
H2S04 acid content, 100% of ulid weight (73-100 gpl H2$04, stoichiometric
requirement for a typical test is a70 gpl).
A large number of leach tests, both in a kettle reactor system and on a
larger scale confirm that sulfuric acid extractions are excellent. Typical
leach -esults are presented in Table 6.8.
The residue from the leach test does not pass the EP toxicity test; Table
6.9. This is a preliminary conclusion that needs to be verified or disproven
during a pilot scale study. The preliminary conclusion is based on CP test
results on three design matrix test residues.
The weight cf residue remaining after a typical leach test is
approximately fifteen percent of the starting solids. The final leach residue
is made up primarily of very finely divided iron and silica bearing compounds.
Example compositions are presented in Section 8.2.
6.3.1.2. Large Scale Leach
The large scale leach testwork produced a concentrated leach solution,
e.g., leach of Barrel one sludge produced 30 liters containing (in gpl): 11.16
Cu, 20.47 Fe. 18.04 Zn. 1.76 Cr, 7.96 Ni, 1.14 Cri, 4.61 Al; leach of Barrel Ib
0
sludae produced 212 liters containing (in gpl): 3.25 Cu, 9.73 Fe. 5.27 Zn,
3.92 Cr. 1.21 Ni, 0.08 Cd, and 1.74 Al. Note the difference between Barrel one
and 18 concentrations. The test assembly is limited by tne iron and zinc
86
-------
TABLE 6.7. STARTING SLUDGE MATERIAL BLENDED SAMPLE REPRODUCIBILITV
Test No.
227
228
229
Average
2486
2487
24B8
2489
2490
Average
Compost t ton tn Solid
Cu
2.41
2.41
2.48
S.lliO.O)
8.26
6.57
F.03
4.41
4.05
4.86,2.36
-fi-
ll. 33
11. d8
11.65
ll.«?,0./9
19.05
18.06
17.37
17.16
16.91
ll.llll.14
Zn
8.40
8.45
8.75
a.SliO.22
6.15
8.61
9.99
9.83
6.96
e.iuz.ie
Cr
Barrel
1.36
1.35
1.35
I.»t0.0l
Barrel
B.52
7.10
6.24
4.46
4.13
t.uS.2 41
HI
5 (used
5.08
4.08
5.08
4.99iD.9l
Al
In kettle
4.05
4.15
4.55
t.IfeO.IO
18 (used tn large
1.91
2.23
2.2(1
2.45
2.47
2.27j0.1S
2.66
2.81
3.13
2.77
2.58
:.»i0.i4
(X)
Cd
test)
0.39
0.41
0.41
O.tOiO
scale
....
0.04
0.08
0.12
0.11
0.09iO.
Ca
1.08
1.00
1.10
ei i.odo.04
test)
0.3!
0.45
0.44
0.64
0.64
01 OMif.lt
Pb
0.09
0.10
0.09
0.09iO.OI
0.08
0.08
0.07
0.05
0.11
0.01. .01
Ha
0.68
0.57
0.76
O.tftiO.IO
0.57
0.52
0.65
0.53
0.41
O.&t^O.II
P
.
.
.
4.01
3.78
3.19
2.M
2.54
1.24.0. »
-------
TABLE 6.3. TYPICAL SULFUR1C ACIO
Test No
535
942
532
2492
Condition
100 gm sludge
650 gm sludge
1.000 gm sludge
50,600 gm sludge
Fe
92.0
95.4
55.8
92.0
LEACH OF MIXED METAL HYDROXIDE SLUDGE:
Cu
93.7
94.9
94.3
93.7
Metal
Zn
95.9
90.5
94.2
95.)
Extracted
Hi
95.9
97.8
85.0
9S.9
(X)
Cr
96.5
96.7
96.7
96.5
STANDARD CONDITIONS
Cd
93.0
100.0
97.0
93.0
Al
89.9
95.7
96.0
96.9
Notes: . Standard conditions: one-half hour leach; ambient temperature; sludge/liquid ratio * 0.8;
acid content equivalent to weight of solids in sludge.
. Detailed experimental results presented In Sections 8.2 and 8.13.
83
-------
TABLE 6.9. I.P. TOXIC ITV PROCEDURE APPLIED TO LEACH RESIDUES: EXPERIMENTAL RESULTS
Sample
370
371
372
373
C.P. Leach Procedure Results (rog/H
Cu
5.88
(5.91)
9.00
(8.99)
12.04
(12.39)
24.51
(24.4)
Fe
<0.006
(<0.006)
6.65
(3.92)
«O.OOG
(<0.006)
<0.006
(<0 006)
Cr
2.17
(2.17)
1.81
(2.02)
2.18
(2.28)
0.39
(0.42)
Hi
59.8
(56.8)
151.5
(146.0)
92.42
(93.31)
129.5
(134.7)
Zn
412.1
(388. 3)
401.9
(412.1)
132.4
(137.1)
648.1
(650.8)
Cd
10.17
(8.91)
12.48
(11.65)
6.27
(5.19)
94.06
(97.0)
St
< 0.006
« 0.006)
< 0.006
< 0.006)
< 0.006
< 0.006)
< 0.006
< 0.006)
Al
2.13
(3.27)
25.10
(23.84)
4.28
(3.01)
0.60
(1.41)
Ca
?9.4
(2R.4)
43.0
(42.8)
22.62
(23.22)
585.1
4584.1)
P
< 0.076
U0.076)
2.58
(2.95)
10.31
(3.79)
)4.4
(19.6)
Pb
1.31
(1.30)
2.53
(3.24)
«D.L.
(«D.L.)
1.76
(atrlx test 291; residue 372 resulted from matrix test 356.
-------
content in its ability to treat solutions, i.e., the jarosite precipitation
unit operation cannot be used to effectively treat iron contents above about
10-15 gpl; the zinc solvent extraction unit operation cannot be used to
effectively treat zinc contents above 5-6 gpl. Therefore, in some cases the
solid/liquid ratio used in the leach was varied to produce the desired solution
composition (of iron and zinc) or alternately concentrated leach solutions were
diluted to achieve desired solution composition.
A number of large scale leach tests have been performed. The results are
reported in the sequential data tabulation presented 1n Appendix 8.13, Tables
8.121-8.126. A summary of the large scale testwork is presented 1n Table 6.10.
The extraction results are excellent and comparable to the results obtained in
small scale testwork.
The large scale leach operation (75-100 pounds of sludge per day) can be
accomplished in a single vessel in one-half hour reaction tine; then the
conditions changed to favor iron removal by jarosite precipitation and the iron
removal operation performed in the same vessel. An alternate approach ii to
leach continuously in a much smaller reactor and store the solution for later
jarosite treatment.
The leach residue has poor settling and filtration properties. The
residue blinds the filtering media, i.e., filter papers (small tests) or filcer
cloths (large tests). The poor fi'terability of the leach residue was a major
raason for adopting a treatment procedure based on precipitation of jarosite
into the leach residue. This greatly enhances the solid-liquid separation
process. A comparison of filterability between leach residue and leach
residue-jarosite mixtures in the pilot scale filter press Is presented In
Appendix 8.5, Tables 8.63 and 8.64. Rates are tremendously different, e.g.,
leach residue, 4.5 kg/m /hr., leach residue-jarosite, 25-55 kg/m /hr. Host of
the large scale sol id/11 quid separation testwork was, therefore, conducted on
leach residue-jarosite mixture?. (This aspect of the study is discussed in
more detail in Section 6.3.3 and Appendix 8.5.)
90
-------
TA3LE 6.10. SUMMARY OF LARGE SCALE LEACH TCSWORK
lest Designation __ I Extracted
Fe Cu 2n N1 Cr M Al
w-4 M-6 90-° 90-3
95-9 9*-5 93-° »•»
S1s 6S-° ff-° 96-9
NOTES: Standard IfoSO* leach conditions used for each test except sequential test series five. See
Tables 8.126-8.127 for detailed results.
-------
6.3.1.3. Large Scale Leach (Phase II)
The large scale leach on Phase U material was conducted in the same
manner as In Phase I. The resulting ieach solutions were high chromium, nickel
bearing solutions. Table 6.11. these solutions were then doped (after solids
removal) to achieve desired iron, copper and zinc contents for subsequent
testwork.
The leach soli as were removed from the solution using the filter press.
Poor filterability of the leach residue was overcome by use of Udylite Oxyfiii
985 filter aid. Filter rates comparable to jarosite filtration was achieved hy
use of 9.2 grams of filter aid per square decimeter of filter area. Detailed
experimental data and discussion of results are presented in Appendix Section
8.5.
6.3.2. Iron Removal
€.3.2.1. Iron Removal from High Iron Bearing Solutions
Iron must be removed early in the treatment sequence because of its
coextraction and therefore contamination of subsequent metil separations.
However, alternates do exist as to where in the treatment seojence it is
removed, e.g., iron can be removed prior to any other metal ion by jarositc
precipitation from an acid solution or iron can be removed after copper
extraction because the commercial reagents available for copper extraction are
highly selective for copper over iron. The advantage of removing iron by
jarosite precipitation prior to cooper extraction is that the jarosite
precipitation conditions appear to significantly improve the copper solvent
extraction process phase separations. Shake tests for copper extraction (using
conmercial reagents) applied to untreated leach solutions produce system muck
that hinders the separation of the organic and aqueous phases. However, shake
tests (and large scale tests also) show nucn improveo pnase separation, after
iron removal, I.e., the high temperature jarosite precipitation process
produces a leach solution much staoier to treat for copper extraction.
92
-------
TABLE
Sample No.
3208
3255
3459
3 '.82
3542
3606
3619
3670
6.11. EXAMPLE LEACH SOLUTION COMPOSITIONS FOR PHASE
11
MATERIALS
Concentration (gpl)
^^H
2.
3.
2.
1.
2.
2.
Cu__
-
750
130
600
035
04b
225
^v
1
1
3
4
4
3
3
3
Fe
^^•^••^•^v
.728
.611
.899
.068
.237
.616
.117
.078
Zn
2.425
2.231
-
0.131
0.300
0.126
0.099
0.105
6.471
5.470
1.987
2.084
2.238
1.907
1.725
1.693
Nl
2.502
2.547
5.847
6.260
6.415
5.727
5.019
5.046
At
0.029
0.035
0.207
0.248
0.353
0.362
0.354
0.373
i^V
0.
0.
0.
0.
0.
0.
0.
0.
702
696
317
331
369
322
303
276
P
-
0.572
0.656
0.627
-
-
-
Notes: Standard Condittons: One-half hour leach; 40-S5°C; sludge/liquid • 0.8;
actd content equal to weight of solid In sludge.
-------
Iron can be removed by other alternate treatment processes. Such
alternatives are discussed in Sections 8.3.2 and 8.3.4. The emphasis of this
study was placed on removing iron, prior to any other metal, by selecti\c
precipitation as a jarosite compound. The detailed experimental results are
presented in Section 8.3.1. Jarosite precipitation is a rather widely used
commercial means of rejecting iron from an acid leach solution' •'.
There are presently 16 commercial zinc plants using a jarosite
precipitation process' '). All of these plants use either sodium or ammonium
as the alkali ions. Potassium is used in several industrial treatment
flowsheets; usually for those flowsheets that deal with the recovery of a high
value product such as copper*8*9' and cobalt* '.
The extent of iron removal from a solution is system dependent. However,
some generalities can be stated that assist in the design of an appropriate
Iron renewal system, i.e., Dutrizac* ' has recently reviewed and summarized a
great deal of literature on jarosite precipitation studies. The results of a
portion of his review on the conditions affecting the precipitation of jarosite
family compounds is paraphrased below:
'Sodium, Potassium and Ammonium Jarosites
Jarosites for each alkali cation exist and can be readily formed.
Most research has been performed on sodium and amironium Jarosites
because of the iower reagent cost. A substantial body of research
information exists for the jarosite families.
'Temperature
Jarosites can be formed at room temperature but the rate of
formation is very slow, e.g., potassium jarosite was formed at 25 C
in a pH range of 0.82-1.72 but required four weeks to six months.
Jarosite precipitation is quite rapid above 80 C. Commercially
useful rates require temperatures greater than 90 C and sodium and
aomonium require higher temperatures than potassium.
'£H
Solution pH.is very important in the jarosite precipitation process,
i.e., the precipitation reactions produce acid and if the pH is not
controlled the reaction is stopped. For example the jarosite
precipitation reaction produces one mole of acid for each mole of
94
-------
Fe precipitated. Outrizac presented the results of Babcan^11' for
the pH-temperature stability of potassium jarosite. Figure 6.4. The
pH range at which jarosite forms decreases in maximum value as the
temperature is raised, i.e., at 20 C the range is 2-3, at 100 C it
1s 1-2.3. The present work was conducted in the range of 88-92°C at
pH values of 2-2.7. A summary of the research results of several
workers seems to suggest that the ideal range 1s 1.5-1.6 at 100 C.
too
§'»
2 100
40
»
0
FtO-OH
0 I
to it it a
Figure 6.4. Stability field for potassium jarosite formation (hatched
area) as a function of pH and temperature for jarosite
formation from 0.5 M Fe2(S04)3 solutions at 20-200 C.
The critical pH at 90°C at which the crystallizing
(excellent filtering) transforms to an amorphous form (poor
filtering) is
pH
0.21 log [Fe*3]
*3
(0 grams per liter zinc)
(100 grams per liter zinc)
1.84
pH - 0.21 log [Fo] + 1.80
where [Fe ] - grams per liter
'Alkali Concentration
The removal of Fe* appears to be essentially independent of alkali
concentration. Slight excess stoichiometric amounts seem
appropriate.
'Iron Concentration
Jarosltes are readily precipitated from solutions containing
0.025-3.0 M Fe •* (1.3-167.4 gpl). The lower limit for Fe J
concentration appear* to be about 0.001 M ( 50 ppm) .
-------
'Order of Stability
The extent of iron precipitation is in the order K > NH. -vNa.
Therefore, lower solution iron contents may.be expected using K
ions. The free energy of formation values* ' fo«- the jarosites
are: -788.6 Kcal/mole. -778.4 Kcal/mole. -736.2 Kcal/mole for K ,
Na , NH4" respectively.
'Ionic Strength
Studies on formation of potassium jarosites from high ionic
strength solutions show no appreciaole effect.
'Seeding
The precipitation process is dependent on the presence of a seed.
Host investigators suggest approximately 100 gpl seed. Recycling
of seed is recommended so that large crystalline jarosite particles
form in order to enhance the settling or filtration rate. In the
present project the leach residue serves the function of a seed
nucleation site.
'Final Iron Content Achievable
Industrially the jarosite process is used to decrease the iron
content from very high levels to 1-2 gpl. The equilibria
relationship for ammonium jarosit^.at 100 C shows that very low
Iron content should be achievable* ':
3 » 0.004 (gpl)
(121
Plant practice shows equilibrium is not truly attained* ' and the
relationship is:
CFe ] = 0.01 (gpl)
[H2S04]
Therefore, low iron concentration is possible, e.g., at pH » 1 the [H9SOA]
- 4.9 gpl and [Fe °] - 0.049 gpl or < 50 ppm. c *
The iron level achieved in the final solution depends on time, pH,
temperature, and alkali ion used. The iron contents achieved in
this study for large scale testwork usually was in the range of a
few hundred ijg/liter for the conditions: pH = 1.8-2.5, temperature
o 88-92 C, K alkaline ion, time 5-6 hours. Iron contents in the
96 .
-------
range of ?00-500 mg/liter are considered appropriately low for
subsequent zinc solvent extraction.
'Impurity Behavior
The partitioning of impurities to the jarosite product has been
considered. The following generalities are noted. The extent of
incorporation of impurities in the jarosite solid product
Increases:
'with Increasing M S04 concentration in solution
'with increasing pH
'with increasing alkali concentration in solution
'with decreasing Fe concentration in solution
*K jarosite > Na jarositetNH4 jarosite.
The pcder of-cation^metal.incorporation appears to be Fe » Cu*2
> In2 > Co*Z > Mi *' > Cd*S
A table describing the partitioning of impurities* ' between
potassium jarosite and the solution is presented in Table 6.12.
The K value is defined by the ratio:
height percent in jarosite
concentration of- impurity in solution (g/100 cc)
Saarinen* ' investigated the incorporation of Cr , Co and Ni from
solutions (the concentration levels not given) into sodium jarosite. His
+2
results showed Incorporation to be low: 0.3*1.4 wt. X Ni ; 0.5-1.4 wt. J
Co*2; and 0.6-1.6 wt. 1 Cr*3 at 90°C in the pH range 1-2.
Some anions also are incorporated into the jarosite structure. Chromate
may substitute completely for sulfate in jarosite compounds(3,4,15). Some
anions co-precipitate with rather than incorporate.in the jarosite structure.
Dutrizac* ' has summarized the results of a number of studies dealing with
anion behavior during jarosite precipitation. A portion of his results are
presented in Table 6.13.
97
-------
TABLE 6.1Z. RELATIVE PARTITIONING OF SOME IMPURITIES BETWEEN
POTASSIUM JAROSITE AND THE SOLUTION PHASE 03)
Impurity
Zn+2
Cd*
Cu+2
Mg+2
N1+2
Al+3
Initial Concentration (g/lOOcc) K
3.2
O.C56
0.32
0.78
- 16.3
- 1.2
- 1.60
- 1.65
0.0015
0.63
- 1.4
0.20 - 0.08
0.02
1.0 - 0.56
0.013 - 0.006
0.7
1.4 - 1.3
TABLE 6.13. BEHAVIOR OF SOME ANIONS DURING JAROSITE PRECIPITATION 0)
Ionic Specie Precipitation Behavior
Cr04~2 Substitutes for Sulfate in jarosite
structure, KFe3 (Cr04-z)2(OH)6
Mn04~ (>0.05 M) Precipitates with ferric ion as
poorly crystalline (Mn.
(<0.03 M) Precipitates with jarosite as poorly
crystalline (Mn, Fe)02
Si*4 (>0.05 M) Forms silica gel at 97°C
(>0.03 M) Does not precipitate
2 ( 0.25-0.1 M) Hydrolyzes to amorphous (Sn. Fe)02
2 Precipitates with ferric ion as
Fe2(Mo04)3
98
-------
Gordon and Steintveit* ' have reviewed possible jarosite disposal
techniques. A good deal of effort has been expanded to determine appropriate
disposal techniques because jarosites in most cases have sufficient heavy metal
fun content to require use of impermeable membrane lined storage areas.
Treatment processes investigated include:
'Sulfation roasting to solubilize heavy metal salts and to produce
pure iron oxide(16).
"Thermal decomposition and production of iron ox1de(17-21).
'Hydrothermal decomposition to form Iron oxide and recover soluble
saHs(22,23).
'Electric furnace smelting(24).
'Fertilizer(25), especially NH4 and K jarosites.
'Fillers In asphalt or as an iron source In cement(26).
The results of the present study confirm the impoundment or burial of the
jarosite product after thermal or air drying would be necessary. The EP test
shows that the leach residue-jarosite product does release some heavy metals;
Table 6.14. The quantity of leach residue-jarosite solid material produced
.by a waste treatment plant would, of course, be very dependent on the incoming
Iron content. The sludges studies in the first phase of this work were high
In iron, 10-15 Ut. % of the solids. Much of the nation's mixed metal sludges
are, however, not high, in iron; usually less than 2 wt. % of the solids. The
sludges studied in the second phase of the present work were low in iron,
<4 wt. %.
Even if the leach residue-jarosite solids were considered hazardous, the
quantity of material to be disposed of would be considerably less than the
starting sludge material; e.g., the low iron sludges (<2 wt. 2 Fe) would produce
approximately 0.02 gm jarosite sol id/gin sludge (40 pounds/ton sludge); the high
Iron sludges (15 wt. Z Fe) would produce approximately 0.15 gm jarosiwe solid/gm
sludge (300 pounds/ton sludge).
99
-------
IARLE 6.14. E.P- TOXIC11V PROCEDURE APPLIED TO LEACH RESIOUE-JAROSITE SOLIDS PRODUCED FROM FULL SCALE (50.4 kg)
TEST.
Sanple
2612
2613
2614
2615
2616
2617
E.P. Leach Procedure Results (019/1)
Cu
21.5
21.5
21.5
21.4
21.3
21.3
Fe
0.17
0.12
0.11
0.25
0.16
0.24
Zn
7.94
8.90
8.55
9.79
9.51
9.16
Cr
0.33
0.31
0.20
0.28
0.24
0.26
Ni
2.35
2.58
2.60
2.41
2.31
2.23
Cd
0.03
0.03
0.03
0.09
0.04
0.04
A1
0.01
0.01
0.01
0.01
0.01
0.01
Pb
0.27
0.23
0.35
0.24
0.15
0.19
Ca
19.2
18.4
18.9
21.4
20.9
20.9
P
11.0
13.7
13.8
12.1
10.9
10.7
g
'Notes: . Tests performed according to EPA designated EP Toxlclty test*2" .
. Starting Leach residue-JorosHe composition (S): Fe Cu Zn Cr HI Cd Al Pb Ca _P
l».3 :.3 Q.?t 3.20 1.20 0.10 1.70 0.10 0.23 3.19
. Large scale test performed on Barrel IB sludge.
•
. EPA designated concentration of contaminants for characteristic toxicIty (ng/l): Cd Cr _Pb_
1.0 5.0 5.0
-------
6.3.2.2. Iron Removal from Low Iron Bearing Solutions
Iron removal from low iron bearing solutions is difficult by jarosite
precipitation. An alternative method of removal, by solvent extraction, was
investigated. This unit operation depicted in the low iron flowsheet, Figure
6.3, follows copper solvent extraction vhereas jarosite precipitation used on
the high iron solution is conducted prior to copper solvent extraction.
The removal of iron from solutions containing a few grams per liter Iron
by solvent extraction using D-EHPA-kerosene mixtures was experimentally
investigated during the Phase II study. The envisioned advantages included:
no loss of chromium during jarosite precipitation because of the removal of
this unit operation and the generation of a smaller quantity of disposable
solid residue.
• OpEHPA reagent has a selectivity for metal cations that is a strong
function of pH, Figures S.lOa and 8.10b. At low pH levels iron is selectively
extracted from an aqueous phase into the organic phase. A portion of the zinc
is coextracted with the iron al a pH of about cne but this can be effectively
recovered selectively from the organic phase by sulfuric acid (200 gpl)
stripping. Ferric ions are strongly bound within the organic pha.se and are not
stripped easily. Sulfuric acid (at 200 gpl) will not strip the iron.
Hydrochloric acid (4-5 N) will strip the ferric ions.
Experimental studies show excellent iron removal from leach solutions,
e.g., iron contents were consistently lowered to <50 ppm. Other metal cations,
+2 +3
except for zinc, are not co-extracted (Ni , Cr ). The process looks
favorable except for the fact that large quantities of hydrochloric acid is
consumed. Therefore, recovery of hydrochloric acid would be required. There
is one commercial operation that uses D-EHPA loading and hydrochloric acid
stripping with recovery of HC1 by solvent extraction with Amberlite LA-2
(R2NH2C1){31).
Detailed experimental data and discussion of results for the solvent
extraction of iron are presented in Appendix Section 8.5.
101
-------
6.3.3. Solid/Liquid Separation
The separation of leach residue-jarosite solids from the solution phase is
very effectively accomplished by settling and filtration. The major advantage
of the jarosite process is the rapid filtration rate achievable. Industrial
> rar
are achieved* '.
2 2
rates in the range 4,540-13,620 kgm residue/in day '5-15 tons residue m day)
A description of the large scale filter press is presented In Section 8.4.
The procedure used for solids separation was to leach the sludge, precipitate
the jarosite into the leach re'idue, pump the slurry to a storage settler,
allow the jarosite residue to settle, pump most of the solution off the settled
solids, then to filter the remaining slurry through the filter press. The
filter press has several unique features that allow the user to exercise a
variety of operating-cake treatment options, e.g., top or side cake washing,
cake compression, cake drying. Tests were not conducted to determine the best
set of conditions for cake clean-up and recovery during the Phase I study
because of the limited number of large scale tests performed. Even without
optimizing the operating parameters cakes containing only thirty percent
moisture were produced.
The operation of the filter press is straight forward and major problems
were not encountered in the present study. It should be noted, however, that a
coarse screen must be mounted to cover the slurr.v pump inlet hose to prevent
small pieces of wood, glass, pebbles, etc., from entering the diaphragm pump.
6.3.4. Copper Solvent Extraction
Commercial copper solvent extraction processes are described in detail in
a recent publication^28'. The Handbook of Solvent Extraction (1983). A
significant portion of the World's copper is produced by solvent extraction,
e.g., world copper production capacity by SX is over 500,000 tons/year. The
technology is, therefore, well developed.
102
-------
. Several reagents are used commercially but all belong to the hydroxyoximes
family: A comparative summary of the three major reagents is presented in
Table 6.15.
TABLE 6.15. COPPER SOLVENT EXTRACTION COMMERCIAL REAGENTS
129)
Trade Name
LIX 64N
(Henkel Corp.)
P 5100
(Acorga, Ltd.)
LrX-622
(Henkel Corp.)
Composition
2-HydH)xy-5-nonyl
benzophenone oxime
(LIX-65N) plus 5,8-
01 ethyl-7-hydroxy-6-
dodecanone oxitne
Substituted salicyl
aldoxlme plus eoual
amount of nonyl phenol
Not reported in
literature
Comments
Standard Reage.it for
low copper bearing
sulfate aqueous solutions;
usually applied to solu-
tions with <2 gpl Cu
Strong cnelatlng agent.
Useful for high copper
bearing sulfate solutions
Strong chelatlng agent.
Useful for high copper
bearing sulfate solutions.
Requires high (160-200
gpl) acid for stripping
Commerci'al copper solvent extraction has been applied primarily to dump
and heap leach operations. Very little study has been devoted to its use on
complex metal bearing solutions. The dump and heap leach processes are
primarily iron and copper bearing sulfuric acid solutions whereas the sludge
leach '.olutions contain copper, iron, nickel, zinc, chromium and sometimes
aluminum, calcium and cadmium. It was, therefore, of interest and necessary to
Investigate whether SX would be selective toward copper over the other metal
1on constitutents. The equilibrium distribution diagram^ ' shows (Figure 6.5)
that a pH can be selected at which copper should be extracted in preference to
the other metals present.
The experimental procedure used in solvent extraction testwork is
presented in Section 5.2.1. Small-scale shake tests were performed to
determine appropriate experimental conditions for subsequent small-scale
103
-------
s
75
Cu2<
Ni2
6
PH
Figure 6.5. Equilibrium distribution diagram for LIX 64N.
(From Kordosky(29))
104
-------
continuous testwork and ultimately full-scale continuous testwork. The
detailed experimental resu'ts are-presented in Section 8.6, 8.8, and 3.13.
Small-scale testwork showed good selectivity and excellent phase
separation in accordance with quoted literature conditions. Testwork was
performed using LIX-64N , LIX-622, and ACORGA 5100. All appear appropriate for
application to sludge leach solutions. Small-scale continuous testwork was
performed using LIX-64N and LIX-622. LIX-622 was chosen for large-scale
continuous testwork because of its high copper loading capabilities, its
Insensitivity {• jood phase separation) to system temperature, and its fast
loading and stripping capabilities.
Small scale continuous testwork on high iron containing solutions (>15 gpl
Fe) showed a much (see list of definitions, p. xxvi of this report) formation
problem especially when aged (several weeks) solutions were used. Therefore,
most of the subsequent research was performed on jarosited solutions. The
jarosited solutions even when aged did not show a muck formation problem.
Large-scale testwork was performed in a Reister testrack (describee and
shown schematically in Section 5.2.1.2. and pictorally in Section 8.14). The
experimental results for Phase I testwork for five large scale sequential tests
are presented in Section 8.13. The test results for a five-day large scale
test conducted during Phase II are presented in Section 8.6.2. An eleven day
continuous copper extraction and organic degradation test study was conducted
and the results are presented in Section 8.6.3. The copper extraction results
for all the testwork are summarized in Tables 6.16 and 6.17. Degradation test
studies were conducted in the large scale test system and In the bell
Engineering testrack. The important considerations with request to organic
reagent degradation are: the amount of aqueous phase that contacts the
organic, effect.of mixer action on stability or organic reagent to oxidation;
and the effectiveness of the organic to function well over a large number of
load/strip cycles.
105
-------
TABLE 6.16. SUWARY CF URGE SCALE TESTS ON SOLVENT EJECTION OF
COPPER WITH LJX 622
Sample No. Condition Coppsr Extraction From Leach Solution
Percent Copper Content In Solution
Initial (gpl) Final (qplj
Sequential Series One (Table 8.«6)
1524 Rafflnate From 98.9 1.37 0.017
Contact.(40 lit.)
Ser*nt1al Series Two (Table 8.88)
1816 Rafflnate From 94.4 0.39 0.022
Cuitact. (60 lit.)
Sequential Pertes Three (Table 8.89}
2005 Rafflnate From 98.0 2.32 0.047
Contact. (20 lit.)
Sequential Ser1»s Four ("able 8.90)
2146 RaffInate Fro* 96.9 3.89 0.120
Contact. (90 lit.)
Sequertlal Series Five (Table 8.91)
2146 Rafflnate Free: 99.0 3.05 0.030
Contact. (160 lit.)
Note: . Detailed results presented In Section 8.13.
106
-------
TABLE 6.17. SUMMARY Of CONTINUOUS COPPER EXTRACTION: ELEVEN DAY LONG TERM ORGANIC
EXPOSURE TEST RESULTS.
Sample No.
Condition
Copper
Extraction From
Copper Concentration (gpl)
3458
3474
3482
3493
35C1-R
3509
3519
3533
3542
3548
3b52
3567
3C06
3613
3619
3631
Starting Solution
First Day Raffinate
Starting Solution
Second Day Pjff.
Starting Solution
Third D«
-------
CO
TABLt 6.17. CONTINUED
Sample
No. Condition
Copper
extraction From Leach Solution
Copper Concentration (gpl) Copper Extracted (X)
3639
3643
3657
3664
3670
3703
Note:
Starting Solution
Ninth Day Raff.
Starting Solution
Tenth Day Raff.
Starting Solution
Eleventh Day Raff.
. Test conditions detailed
Initial
1.812
2.026
2.225
in Table 8.81
Final
0.073 97.0
0.043 96.0
0.049 97.8
•
-------
There was no apparent effect of continuous exposure of the recycled
organic phase in the large scale system over \8 hours of cumulative exposure,
Table 6.18; 14 1/2 liters of organic were exposed (two stages of load, two
stages of strip) to 274 liters of aqueous leach solution. Therefore, the
aqueous/organic contact ratio was 18.5.
A second series of continuous exposure tests was performed in the Bell
Engineering testrack. There was no apparent effect of continuous exposure over
112.8 hours of cumulative exposure. Tables 6.19, 6.20. Three and eight-tenth;
liters of organic were exposed (three stages of load, two stages of strip) to
341.5 liters of aqueous leach solution. Therefore, the aqueous/organic contact
ratio was 88 (approximately 226 load/strip cycles).
Detailed experimental results and further discussion are presented in
Anpendix Section 8.8.1.
6.3.5. Zinc Solvent Extraction
Commercial application of solvent extraction for zinc recovery is limited.
However, for the treatment of solutions containing a mixture of zinc, chromium,
and nickel the alternatives are few. The only large scale commercial
application of zinc solvent extraction at present is in Spain; Technical
Reunidos uses such a process at ics Bilbao plant for the production of 8,000
tons/yr. of zinc. (Thorsen* ' discusses the commercial operation at Bilbao.)
The commercial reagent available for extraction of zinc (and cadmium) from
acid solutions is the organo phosphoric acid; diethylhexylphosphoric acid
(D-EHPA). The equilibrium distribution diagram^30^ illustrating zinc and other
metal extraction as a function of pH is presented in Section 3.7. Zinc can be
selectively extracted from an. acid solution of pH^2 in the presence of
chromium and nickel. Aluminum and calcium are not presented OP the referenced
distribution diagram. However, if these ions are present in the leach solution
they will be partially co-extracted. Conditions can. in fact, be chosen so
that zinc, aluminum and calcium are completely coextracted.
109
-------
TABLE
Sample
3271
3286
3287
328B
3289
3307
3338
3309
3310
3339
3340
3341
3342
3361
336?
3371
3372
6.18. LIX 622 ORGANIC EXPOSED
No. Organic Exposure
To Aqueous Phase
Starting Aqueous Solution,
First Day
65 liters
• II
None
n
Second Da/
141 liters
H N
None
H
Third Day
209 liters
P N
None
H
Fourth Day_
274 liters
• H
None
None
IN LARGE
SCALE TESTRACK FOR
FIVE DAY TEST PERIOD.
Contacts Copper Concentration in Aqjeous Phase (gpl)
3.12 gpl
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
System Organic
Cu
0.06
0.01
0.08
0.01
0.10
0.02
0.13
0.01
New Organic
0.12
< 0.01
0.07
< 0.00
0.09
< 0.01
0.08
< 0.00
-------
TABLE 6.18. CONTINUED
Sample No.
^
Organic Exp> .re
To Aqueous 1 jse
Fifth Day
347 liters
• N
Contacts
First
Second
Copper Concentatton in Aqueous Phase
System Organic New Organic
0.08
0.01
(gp»
Notes: . ISv/o LIX 622 in Kernac 4708 Kerosene.
. lOOcc of used organic stripped twice with lOOcc clean 200 gpl HjSOo,.
. Stripped system organic contacted with lOOcc of No. 3271 aqueous at initial pll -1.54,
7 mtnutes. 25°C.
. New organic treated sane as system organic but not exposed to leach solution before
test.
. New organic treated with 30 gpl Cu. 200 gpl 112804 before use.
-------
TABLE 6.19. LIX 622 LONG TERM EXPOSURE DEGRADATION TEST SUCURY
Sample No.
Organic Exposure
To Aqueous Phase
Contacts
Starting Aqueous Solution, 3.112 gpl
31711
3479
3400
3481
3495
3496
3497
3498
3514
3515
3516
3517
3536
3537
3510
3S41
First Day
46. s liters
« •
None
•
Second Day
86.5 liters
N «
None
•
Third Day
125.5 liters
It
None
N
Fourth Day
161.5 liters
H H
• None
•
First
Second
First
Second
First
Second
First
Second
First'
Second
First
Second
First
Second
First
Second
Copper Concentration
System Organic
Cu
0.061
0.008
0.103
0.027
0.114
0.016
0.109
0.019
In Aqueous Phase (gpl)
New Organic
0.121
0.006
0.031
0.000
0.035
0.016
0.028
0.001
-------
C.I
TABLE 6.19. CONTINUED
Sample
3550
3551
3615
3616
3617
3618
3635
3636
3637
3630
3647
3648
3649
3650
3665
3666
1667
3668
No. Organic Exposure
To Aqueous Phase
First Day
187.0 liters
None
Sixth Day
206.5 liters
N «
None
•
Seventh Day
241.0 liters
M H
None
N
Eighth Day
275.5 liters
0 M
None
M
Ninth Day
287.5 liters
M •
None
M
Contacts
First
First
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
Copper Concentration
System Organic
0.001
0.036
0.008
(MI2
0.022
0.258
0.023
0.120
0.000
In Aqueous Phase (gpl)
New Organic
0.040
0.006
0.007
0.034
0.007
0.053
0.004
O.OIS
0.020
-------
TABLE 6.19. CONTINUED
Notes: . Conditions for each days exposure given In Table 8.02.
. Degradation test conditions: 50cc system organic stripped twice (0/A >1)
with unused 200 gpl H^SO/j; stripped organic contacted with copper stock
solution, pll • 1.36 for first four tests, pll « 2.0 for last five tests; a
second system organic sample contacted same stock solution, i.e., stock
solution was contacted twice with two used organic samples, stock pH not
adjusted between contacts.
Unused organic same as system organ•". 15 x LIX 622, contacted with a 30
gpl Cu. 200 gpl I^SO^ solution, then contacted with stock solution as
described above for system organic.
-------
Sample No.
TABLE 6.20. LIX 622
LOADING.
Organic Exposure
To Aqueous Phase
LONG TERM EXPOSURE DEGRADATION TEST SUMMARY:
Contacts
Loading.
.System Organic
Stock Aqueous Solution. 3.1)2 gpl Cu. 3.958 gpl Fe, 0
2.014 gpl Cr. 6.06) gpl Hi. 0.287 gpl Al. 0.319 gpl
3478
3470
3480
34S1
3495
3496
3497
3496 '
3514
35)5
3516
351 >
3536
3537
3540
3541
First Day
46.5 liters Aqueous
M H
Hone
M
Second Dd£
86.5 liters
H •
None
M
Third Day
125.5 liters
•
None
M
Fourth Daj;
161.5 liters
M
None
N
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
0.203
0.004
0.200
0.005
0.200
0.006
0.200
0.006
gpl/X LIX 622
N»w Organic
.122 gpl Zn.
Ca
0.199
0.008
0.205
0.014
0.205 '
0.001
0.206
0.002
-------
TABLE 6.20. CONTINUED
Sample No.
3550
3551
3615
3616
3617
3618
3635
3636
3637
3638
3647
3648
3649
3650
3665
3666
3667
3668
Organic Exposure
To Aqueous Phase
Fifth Day
187.0 liters •
None
Sixth Day
206. 5
M
None
N
Seventh Day
241.0 liters
•
None
«
Eighth Day
275.5 liters
•1
None
«
Ninth Day
287.5 liters
n
None
•
Contacts
First
First
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
Loading,
System Organic
0.207
0.205
0.002
0.200
0.006
0.190
0.016
0.199
> 0.013
gpl/S LIX 622
New Organic
0.205
0.207
0.206
0.001
0.204
0.003
0.206
Note: . Conditions for each days exposure presented In Table 8.83.
-------
The leach solutions investigated in the current study contained zinc (5-6
gpl), aluminum (2-3 gpl) and small amounts of cadmium (0.2-0.3 gpl), calcium
(0.5 gpl) and iron (0.2-0.5 gpl). The results of detailed experimental studies
are presented and discussed in Section 8.7; large scale Phase I testwork (zinc
removal after jarosite treatment and Cu SX) is presented in Section 8-. 13.
Large scale and continuous Phase II testwork results (iron removal by SX rather
than by jarosite precipitation) are presented in Section 8.4. The experimental
procedure 1s described in Section 5.2.2 and the large scale equipment Is shown
schematically in Section 5.2.2.2, and pictorally 1n Section 8.14.2.
••
6.3.5.1. Large Scale Zinc Solvent Extraction (Phase I)
The results of large scale Phase I testwork are summarized in Table G.21.
Zinc can be Affectively extracted by D^EHPA. An apparent upper limit on the
zinc content in the leach solution is 5-6 gpl using .1 forty volume percent
O.EHPA reagent mixture. Seven cells are required to accomplish effective zinc
recovery from the leach solution; four stages of contact; three stages of
strip. Solution pH adjustment is required after the first two contacts In
order to ensure zinc removal to <70 mg/1.
Iron (Fe ) -is coextracted by OgEHPA. It is not stripped by sulfuric acid
and, therefore, occupies extractant sites in the organic phase. Iron must be
removed from the organic or else it will blind up all the sites over a period
of time. Iron can be stripped from the organic phase by 4-6 N hydrochloric
acid. Therefore, the proposed treatment process consists of: extraction of
zinc, residual iron, calcium and aluminum frc:n the leach solution in four
stages of pH controlled contact; strip of the zinc from the organic by 200 gpl
H-SO. in three stages of contact; removal of an appropriate amount of bleed
solution from the sulfuric acid stripped organic; strip the bleed organic phase
with 4-6N HC1 to remove the Fe*1* and Al ; recycle the bleed organic back to
the system organic phase going into the extraction stages. If calcium is
present in the leach solution it v.-ill be extracted with the zinc and
subsequently will be stripped by sulfuric acid in the strip cells. It forms
gypsum solid that can be continuously filtered from the aqueous strip phase.
117
-------
TABLE 6.21. SUMMARY OF URGE SCALE TESTS ON SOLVENT EXTRACTION OF
ZINC WITH 02EHPA
Sample No. Condition Zinc Extraction From Leach Solution
Percent Zinc Content In Solution
Initial (gpl) Final (gpl)
Sequential Series One (Table 8.86)
1532 Rafflnate From 60.S 5.14 1.00
Contact. (25 ilt.)
Sequential Series Three (Table 8.89)
2109 Rafflnate From 97.4 5.70 0.15
Contact. (20 lit.)
Sequential Series Four (Table 8.90)
2181 Rafflnate From 97.8 5.89 0.13
C.-ntact. (50 lit.)
2256 Rafflnate Froo 98.8 4.94 0.060
Contact. (90 lit.)
Sequential Series Five (Table 8.91)
2526B Rafflnate From 98.9 6.20 0.070
Contact. (160 ll^
Note: . Detailed results presented in Section 8.13.
118
-------
Z1nc solvent extraction appears appropriate for selectively removing zinc from
a chromium and nickel bearing solution; and for eliminating calcium, iron and
aluminum from the leach solution.
6.3.5.2. Zinc Solvent Extraction (Phase II)
As noted previously the major difference In flowsheet testing between
Phase I and Phase II was that iron was removed in Phase I testwork by jarosite
precipitation with residual iron removal concurrent with zinc solvent
extraction. Whereas Iron (present at much lower concentrations) was removed In
Phase II testwork by solvent extraction.
The results of large scale Phase II testwork are summarized in Taole 6.22.
The continuous testwork to determine reagent loss rates and potential
degradation of reagent are summarized in Table 6.23.
The large scale testwork was conducted in a series of ten cells; one cell
for preferentially loading iron; one cell for stripping zinc from the iron
loaded organic; three cells for stripping iron loaded organic; three cells for
zinc loading; and two cells for stripping zinc loaded organic. Zinc is
effectively extracted from the aqueous leach phase, i.e., zinc concentrations
in the aqueous pnase can be lowered tc below SO ppm without appreciable
coextraction of chromium or nickel.
Initially a problem with crud formation was experienced in the iron
extraction cell. (This problem is discussed in greater detail in Section
8.4.3.) The solution to the problem was to use a kerosene solvent containing a
lower aromatic constituent content. A switch from use of KERMAC 470B to KERMAC
510 solved the crud problem.
Organic loss by carryout from the load circuit into the final raffinate
was measured by periodically collecting a liter of raffinate in a graduated
cylinder, allowing the organic phase to separate, then measuring the volume of
organic per liter of raffinate. The carryout rate ranged from 0.25 cc/1 to
119
-------
TABLE 6.22. SUMMARY OF LARGE SCALE TESTS ON SOLVENT EXTRACTION OF ZINC
AND IRON WITH DEHPA (PHASE I!)
Sample No. Condition Extracted From '.each Solution
Content in Solution Percent
InitiaKgpl! Flnal(gpl)
Zn Fe Zn Fe Zn Fe
FIRST DAY.75 lit.
3281-B First Cell Feed 1.8)5 1.164
3284 Raff mate 0.014 D.L. 99.2 100.0
•SECOND PAY.75 lit.
3351 First Cell Feed 2.208 1.532
3328 Raffinate 0.026 0.0*0 98.7 97.4
THIRD DAY.75 lit.
3351 First Cell Feed 2.270 1.H91
3367 Raffinate 0.0*3 0.010 93.1 99.5
FOURTH DAY.75 lit.
3414 First Cell Feed 2.436 2.362
3434 Raffinate O.C61 0.053 97.5 97.8
Notes: . Test conditions presented In Table 8.48.
results presented in Table 3.49.
120
-------
TABLE* 6.23. SUMMARY OF LONG TERM CONTINUOUS TESTUORK: ZINC AND U*ON
REMOVAL
Sample Nc. Condition Extracted Front Leach Solution
Content in Solution Percent
Initlal(gpl) Flnal(gpl)
Zn Fe Zn Fe Zn Fe
FIRST PAY. 19 lit.
3745
3787
3805
3835
3846
3863
3881
3308
3926
3944
3953
3969
3992
4022
4057
4054
First Cell Feed 1.828 2.023
Final Raffinate
SECOND DAY, 19 lit.
First Cell Feed 0.3S4 2.276
Final Raffinate
THI3D DAY. 19" lit.
First Cell Feed 2.207 2.74*
Final Raffinate
FOURTH DAY. 19 lit.
First Cell Feed 2.128 2.035
Final Raffinate
FIFTH DAY, 19 lit.
First Cell Feed 1.999 2.218
Final Raffinate
SIXTH DAY. !9 lit.
First Cell Feed 2.162 2.127
Final Raffinate
SEVENTH DAY. 19 lit.
First Cell Feed 2.084 2.299
Final Raffinate
EIGHTH DAY. 19 lit.
First Cell Feed 1.067 0.582
Final Raffinace
0.080 0.498* 95.6 75.4
0.094 0.070 73.4 96.9
0.03S 0.027 98.4 99.0
0.046 0.319* 98.5 84.3
0.031 0.022 98.4 99.0
0.050 0.238* 97.7 88.8
0.066 0.051 36.8 97.8
0.043 0.028 96.0 95.2
121
-------
TABLE 6.23. CONTINUED
Notes: . * Iron not completely oxidized.
. Test conditions presented in Table 8.54.
. Detailed results presented in Table 3.S3.
122
-------
O.S4 cc/1. These numbers are very dependent on system characteristics.
Commercially entrainments range up to several hundred mg/1.
A series of continuous exposure tests were conducted to provide long term
degradation data. These tests were conducted in the Bell Engineering'testrack.
There was no apparent effect of continuous exposure over 38 hours of cumulative
exposure. Table 6.24; approximately seven and one-half liters of organic was
exposed (one stage of low pH iron loading, three stages of higher pH zinc
loading, three stages of zinc strapping, three stages of iron stripping) to 150
liters of aqueous leach solution. Therefore, the aqueous/organic contact ratio
was 20 (approximately 58 load/strip cycles).
Detailed experimental results and further discussion are presented in
Appendix-Section 8.8.2.
6.3.6. Chromium Oxidation
Selective removal of chromium from a mixed metal solution containing Iron,
copper, zinc, nickel, aluminum does not appear possible without conversion to
+3
an oxidized anionic form. To accomplish the oxidation of Cr requires a
strongly oxidizing environment. This fact means that the oxidation must be
performed after any solvent extraction unit operation because strongly
oxidizing solutions are expected to degrade the organic extracting
reagents* . Therefore, the most appropriate place in tr.e treatment sequence
is after iron, copper and zinc removal. The emphasis, therefore, for this
study was placed on treating chromium and nickel bearing solutions. For
practically ail cases, actual Irach solutions pretreated for iron, copper and
zinc removal were used for the testwork. The solutions considered in the Phase
I study contained approximately 2-6 gpl Cr* , and 2-5 gpl Ni.
Detailed experimental studies are presented and discussed in Section
8.9.1.; large scale testwork is presented in Section 8.13. The experimental
procedure is described in Section 5.3. Large scale chlorine oxidation
equipment is presented pictorially in Section 8.14.2.
123
-------
TABLE 6.24. DEHPA LONG TERN EXPOSURE DEGRADATION TEST
Sample No. Organic Exposure Contacts
To Aqueous Phase
4025
3841
3942
3843
3844
3874
3875
3876
3877
3909
3910
3911
3912
3947
3948
Stock Aqueous Solution,
pH « 2.0. 11.639 gpl Fe.
11.192 gpl Zn
First Day
19 liters aqueous
• •
None
•
Second Day
38 liters
• «
None
•
Third Day
57 liters
m m
None
N
Fourth Day
76 liters
• •
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
Loading, gpl (Zn»Fe)/X DEHPA
System Organic New Organic
0.257
0.068
C.32/
0.032
0.261
0.054
0.270
0.056
i
0.242
0.066
0.286 '
O.OSI
0.271
0.056
-------
in
TABLE 6.24. CCNTINUEi)
Sample No.
3947
3948
3982
3983
3984
398S
4026
4027
4028
4029
Organic Exposure
To Aqueous Phase
None
M
Fifth Day
96 liters
N M
None
N
Sixth Day
115 liters
ii a
None
N
Contacts •
First
Second
First '
Second
First
Second
First
Second
First
Second
Loading, gpl
System Organic
0.248
0.054
0.2P6
0.029
(Zn«Fe)/l DEHPA
New Organic
0.259
0.060
•
0.278
0.06S
Notes: . Detailed experimental results presented in Table 8.95.
-------
A relatively large number of oxidation possibilities were considered.
Only two oxidation techniques are considered feasiole because of reagent cost.
They are chlorine oxidation and electrochemical oxidation.
• *
Large scale testwork using chlorine oxidation showed that slurry oxidation
of precipitated chromlun was effective and controllable. The operational
procedure consisted of adjusting the solution pH to 4-5 (thereby precipitating
most of the Cr*3 as Cr(OH)3); sparging in Cl» gas to raise the solution Eh to
>1,000 mv; and then allowing the reaction to proceed .for several hours. Eighty
to ninety five percent of the chromium was oxidized to the dichromate form.
The unoxidized Cr(OH). solid can either be separated from the solution and
recycled to tne leach stage or left in tne reactor to become a part of the next
oxidation treatment.
The oxidation time period used to treat large volumes was relatively long
in the large scale testwork. However, short time exposures were effective in
the imall-scale testwork. The reason for the difference was the effectiveness
of the contact system used. A more appropriately designed reactor was tested
In the Phase II study, but the results did not show an Improvement in the
oxidation rate. Section 8.9.1.1.2.
Oxidation of chromium achieved by chlorine sparging was In the range
70-80% for a contact period of 4-5 hours at a pH of 4-5; the oxidation achieved
by use of an aspirating chlorinator was in rang? 40-70% for a contact period of
4-5 hours at a pH of 4-5. Complete oxidation Is not, however, required because
the solid residue contains the recxidized chromium as chromium hydroxide. It
1s not lost from the system but is recycled to the initial leach unit
operation. The filtrate solution from the solid/liquid separation (that is
further treated for chromium recovery) is essentially chromium anior.s, e.g.,
99? Cr*6 (as an anion), 1% Cr*3.
Electrochemical oxidation is commercially practiced for regenerating spent
plating baths(32,33). The cell electrodes are separated by a cation permeable
membrane. Chromium is oxidized in the anode compartment and Impurities 1n the
126
-------
anolyte transport across the membrane to the catholyte. This fystem appear-, :o
be applicable to the present treatment sequence; cnromium could be oxidized ana
copper could bo removed from the copper SX strip acid. The operational
procedure would be: pump the leach solution into the anode chamber for
Chromium oxidation; pump the copper bearing strip solution from the copper SX
strip cells into the.- cathode chamber for copper extractfon; recycle the copper
depleted strip solution to the copper SX strip cells to pick up more copper;
pump the oxi-'
-------
TABLE 6.25 SUMMARY OF CONTINUOUS ELECTROCHEMICAL OXIDATION OF
CHROMIUM
Sample No.
5003
5006
5013
5039
Condition
Batch Test
Series One Test
(Table 8.105)
Series Two Test
[Table C.106)
Series Three Test
(Table 8.107}
Chromium Oxidation (I)
85.4
78.0
87.2
95.6
Notes: . Test conditions presented in above referenced tables in
Section 8.9.1.2.
128
-------
the pH range 4-5; lead chromate forms as a dense, crystalline precipitate; the
solids are allowed to settle and the chromium free (<7mg/l Cr) solution is
pumped away from the settled solids; the remaining slurry is recontacted witn
more oxidized chromium bearing solution and the process repeated. The exchange
reaction is complete in a few minutes time. The lead sulfate can be •
regenerated and the chromium recovered in a concentrated form by leaching the
solids in sulfuric acid. The regenerated lead sulfate is then recycled to the
precipitation vessel.
The basis for the precipitation process is shown in Figure 6.6. The
diagram shows that PbCrO. (solid lines on the diagram) Is the stable phase
above pH levels of: ? for SO^/Cr « 1; 2.8 for SO^/Cr « 10; and 3.7 for SO./Cr
• 100. Therefore, the pH range of 4-5 Is appropriate for all the test
conditions used in the present study to produce PbCrO-. It also demonstrates
o
that tJie redissolutlon of PbCrO^ can be achieved at pH levels below the
Intersection between the dashed and solid lines, i.e., for a SO./Cr * 1 a pH <2
will convert the PbCrO. to PbSO..
4 4
The detailed experimental results are presented in Section 8.10.1.1;
large-scale sequential testwork in Section 8.13. The large-scale test results
are summarized in Table 6.26.
6.3.8. Nickel Precipitation
Nickel sulfide can be effectively precipitated from a nicicel bearing
solution by the addition of a sodium sulfide solution. If the sulfide solution
1s added at the proper concentration and rate tt-pre is no release of hydrogen
sulfide gas. The precipitation is performed on the solution resulting from the
removal of chromium, which is at a pH of 4-4.5. Precipitation of nickel from a
solution at this pH value results in effective nickel removal, e.g., large-
scale testwork showed the following results. Nickel was decreased from 2.27
gpl in 3.5 liters of sequentially treated leach solution to 7 mg/liter. Nickel
was decreased from 1.67 gpl in 42 liters of sequentially treated leach solution
to 6 mg/liter.
129
-------
- 4
- 6
- 8
i
ex
*" -10
01
o
•12
-14
CR a 4 GPL,0.077MOLEa/LlT.
|S04-)/(CR) = I
(ay)/(CH) = 10
(so •)/(CR) = 100
HSO
1
1
1
1
01234
Solution pH
Figure 6.6. Lead chranate-lead sulfate stability diagram
Hcno
-------
TABLE 6.26. SUMMARY OK URGE SCALE TESTS ON CHROMIUM PRECIPHATION
Sample No. Condition Chromium Removed From Solution
Percent Oironrfua Content In Soljtlcn
Initial (gpl) Final (gpl)
Sequential Series Four (Table 8.90)
2347 Starting Solution - 1.65
(10 liters)
2376 Final Filtrate 99.5 - 0.008
after 30 mln. exposure
Sequential Series Five (Table 8.91)
2600 Starting Solution - 2.34 •
(42 liters)
2602 Final Filtrate 99._7 - 0.007
after 30 mln. exposure
Hote: . Detailed experimental results presented In Section C.13.
131
-------
The prec'pUatlon is rapid and complete in less than one-half hour.
Therefore, ? relatively small reaction vessel is appropriate for the
precipitation. A deficiency of sodium sulfide should be used so that solution
sulfide ions do not exist. Otherwise hydrogen sulfide .iould be generated when
this solution is recycled to the leach unit operation.
The detailed experimental results are presented in Section 8.11.1.; large
scale testwork 1n Section 8.13. Alternate nickel recovery possibilities are
presented in Section 8.11.2 and 8.15.7.4.
6.4. ECONOMIC ANALYSIS
The following cost summary Is presented as a first order estimate*49'50^
for the flowsheet presented in Figure 6.7. Itemized equipment lists were used
where possible and literature quoted cost figures were used when available.
Costs were estimated using the data of Mular^49^, Word^50', and D>venport*39^.
All costs have been updated to second quarter 1984 using the Marshall and
Stevens (MSS) Index. The current M4S Index value is 794. Mai or cost Items
have been Included. The factored capital cost totals take care of minor
equipment, Instrumentation, processing piping, auxiliary engineering cost, and
plant size factor. Detailed cost sheets, both for capital and operating costs,
are presented in Appendix 8.15.
This is certainly not a detailed engineering cost analysis. It is only
what. Mular and Davenport claim for the calculational technique. I.e., a first
order estimate. Mular suggests that the cost totals will be within +30
percent. If cost were not available for the present flowsheet individual unit
operation capacity but dita existed for a similar commercial unit operation the
sixth tenth rule was used. I.e.,
cost
The potential value of products* ' and reagent costs were obtained from
current literature quotations and are reported in Tables 6.27 ana 6.28.
132
-------
RECYCLE
CJ
SLUDGE
FEEDER
Z.JIph,
II REAMS
|ACIO.«Jsph.l030gp
-------
LEACH COPPER SOLVENT EXTRACTION
SCLI"1?! .1 IOADED ORGANIC
.
I i
H
s
1
0
!
A
*
S
H
4
f»
0
£60 j«l MIXER SEIILCRS
1 ^"ccLtst
STRIPPED ORGANIC | 840 gal K
I 1 1
H
S
1
1
RATH MATE
pll • 1.3-1.4
•
STORAGE
45.000 gal
A. M
""*"
S Pfl
H
i 1
330 gal/d
?rCELLst "" 1
UEPLETtO ELECTROLYTE
JX 6J2
EGNANT El EC
,
'
1— -!-,J
1
IBOgpl II2S04, 30 gpl Cu
I60gpl «2S04. 4} gpl Cu
ELECTRO*INNING
Cul.SOppd
Figure 6.7. Flowsheet for treatment of 50 tpd of mixed metal sludge (continued).
-------
ZINC AtlD IRON SOLVENT EXTRACTION
FKOH S10RAGE
pH:i
M»
IRON LOADED ORGANIC
S
H
IRON DEPLETED ORGANIC ZINC JTRIPPINO (4CELLS)
0
1
S
H
A
t
H
S
~H
0
L
s
H
-
t
N
S
0
1
S
H
A
«*-
f
M
S
4-
0
— J-—
H
IRON STRIPPING
«
^V
f
M
S
4-
1
, b
H
ORGANIC
From » 7.
660 ge MIXER-SETTLER
NoOII
pH - 2
pll AD JUS 1
EXTRACTION
|4 CELLS)
790 gal OEirA
1190 gal KERMAC 910
700 pounds HCI
1/90 pounds H2S04
lOi gal/d NuOII , 900 gpl
ZINC
RAmtlAlE
pll - 1.3
To Cr Oxldatl
*
To Coll l\
> Heed
STORAGE
49.000 gal
CRTSTAL-
LI2ER
ZNSO RECOVERY
ilC I
I!CL STRIP
UUi.
STRIP
SX
IKON
HCL RECOVERY
iOppd FeCI,
Figure 6.7. Flowsheet for treatment of 50 tpd of mixed metal sludge (continued).
-------
FROM STORAGE
— — CHROMIUM OXIDATION— —I
ELE
3000 gal cell
1900 amp
24 hours
CTROCHEMICAL OXIDATION CELLS
TO STC
PbSO .
RAGE
NaOM NaOII |PDSO.( 7430 ppd total)
t ft4
pll:3-4 pll:3-4
|O.M O.jl PRCCIPITATION VESSELS
1040 oal
P« ' >« rJ-D^-i^ «" PbSVsO.. 60DBd
1' •
CHROMIC ACID PRODUCTION
pll-0
| RELEACH VESSEL
J- 100 gal
V"^^
r-L-O^>«»^,
^^"^^ ^^^^_^^«a ^J "^^^*W*
F ORU7I FILTER
H ero 74^i
7900 ppd solution RECYaE 10
230 gpl Cr FKECIPITA1E
VESSEL
Figure 6.7. Flowsheet for treatment of 50 tpd of mixed metal sludge (continued).
-------
NICKEL RECOVERY
tJ
Na.PO
No.
PRECIPITATION VESSELS
1040 gal
TO PONDING
|— NICKEL CONCENTRATION —j
(OR
»vw vrH
sol Id and solullon
Ih
40S|&Ol.
lOOgal
2390 ppd NI,(PVj Mrkot
35ppd
• ii
osKo
, RtlfALH
- I
24 ppd NdOll. 1020 ppd Na S
NICKEL SULFIOC PRECIPITATION
TO PHOSPHATE PRECIPITATION
OftUH FILTtR
1780 ppd HIS
Figure 6.7. Flowsheet for treatment of 50 tpd of mixed metal sludge (continued).
-------
TABLE 6.27. VALUE OF PRODUCTS AND REAGENT COSTS
Product Cost (S/pound)
Cu 0.60
ZnS04-H20 0.20
H2Cr04 1.18
N1 3.45
N1S 1.72
N10 2.60
PbS04 0.85
H2S04 60 (S/ton)
NaOH . • 3CO (S/ton)
Na2S 470 (S/ton)
Cr203 1.90
Sources: October Issue ENJ
Chemical Market Reporter. September 17. 1984.
138
-------
TABLE 6.
Product
1. Cu
2. ZnSO.-H,0
•t •.
3. H2Cr04
4. N1S
5. N1
6. Cr203
7. N10
8. Credit for
•
28. TOTAL PRODUCT VALUE FOR 50 TPO COST ESTIMATE
Quantity (pounds/day)
1130
3180
2310
1780
1150
.1490
1460
Disposal (Si/gallon sludge)*
TOTAL (1.2.3.4.8)
TOTAL (1.2.3.7.8)
TOTAL (1.2,5.6.8)
Potential Value (S/Yr)
223.700
209.900
899.500
1.010.300
1.309.300
934.200
1.252.700
3.300.000
5.643.400
5.885.800
5.977.100
•Disposal costs vary considerably depending on amount of material that
oust be handled.
139
-------
The return on Investment (ROI) was calculated by the equation:
ROI _ (Value of Products-Annualized Cost)(tax rate)(10C)
~- Capital Cost
• •
m
The cost estimate is based on the flowsheet presented in Figure 6.7.
Capital cost and operating cost were estimated. These estimates are presented
In Appendix Section 8.15. Equipment costs were based en cost equations of the
form:
costnow - a(capacUy)b(M4Snow/M4Sthen)
where a, b are constants for a particular piece of equipment. The constants a
and b are provided for a variety of types of equipment by Mular, Woods, and
Davenport (and in some cases on other literature data).
•
An equipment list was prepared for each series of unit operations, the
cost estimated as described above. The Factored Capital Cost was determined by
using the factors as presented in Table 6.29. An annualized cost was then
determined based on a five year period, 12 percent interest rate. An
operational cost for the series of unit operations was established based on
reagent consumption, manpower requirements, maintenance and power consumption.
The results or the calculations are presented in Tables 8.131-8.145.
Operating cost estimates are presented in Table 8.131. The operating cost
estimates do not include personnel other than operational personnel. The
estimates include: unit operations cost; manpower requirements; maintenance
costs; and energy cost.
The ROI calculations were made oased on the following assumptions:
buildings and land are available; tax rate is 50 percent; Interest rate is 12
percent; pay-off period is five years; a credit of one dollar per gallon of
sludge is allowed; and the plant operates for 330 days per year. The quantity
of material to be treated is based on 50 tpd sludge containing 25 weight
percent solids. The solids contain five weight percent of each element, Cu,
140
-------
TABLE 6.29. FACTORED CAPITAL COST ESTIMATE* •
Cost (WS " 794)
1. Purchased Equipment Costs
2. Installed Equipment Costs (1.4 x Item 1)
3. P.-ocess Piping (30X of 2)
4. Instruirentatlon (10X of 2)
5. Auxiliaries (5X of 2)
6. Outside Lires (5X af 2)
7. Total Physical Plant Costs (Sum of 2 through 6).
8. Engineering and Construction (ZOS of 7)
9. Contingencies (1SX of 7)
10. Size Factor (Small Commercial, 10X af 7)
11. TOTAL PLANT FIXED CAPITAL COST (Sum of 7
through 10)
YEARLY COST, Based on 60 month pay-off period, 12X
interest
•Example EatImat* Form
141
-------
Zn, Cr, Ni, seven ana one-half percent Fe; two percent Al and P; and one
percent Ca. Mass oalances were based on actual experimental data generated
during this study. The mass balance results were used to size the necessary
equipment and are reported in Section 8.IS.
The first order estimate return on investment, based on the flowsheet
presented in Figure 6.7, and Tables 6.30, 6.31, is 41+121
R01 - (5.643.400-2 434.100)(j5)) . 4J+
J.oOO.oUU •—
The oxidation unit operation is a major cost in the overall project cost;
both for capital and for operating expense. It is expected that significantly
lower cost would result using newer technology presently neing commercialized.
(A?\
INCO1 ' has developed a technology based on the use of sulfur dioxide and
oxygen that they are commercializing for cyanide destruction. However, their
research results show solution potentials that are sufficiently oxidizing to
oxidize cnromu'm. In tne case where chromium (+3) and nickel (+2) are present
it may be possible at a pH in the L»«ic regions (pH^ 8) to oxidize a slurry of
chromium (+3) hydroxide and nickel (+2) hydroxide to chromate (Cr04°) and solid
nickel (+3) hydroxide. Therefore, a separation between chromium and nickel may
be possible at a much lower cost than either electrochemical oxidation or
chlorine oxidation. This technology was not being used industrially when the
present investigations were begun, therefore, it has not been experimentally
investigated in this study.
Application of the SO.-O, process to chromium oxidation has not been made
and. therefore, costs are not available. However, a first order cost analysis
can be made by assuming the chromium oxidation rate will be similar to the
measured nickel oxidation rate and by costing out the equipment required to
achieve the oxidation. The anticipated unit operations are depicted in Figure
6.8a. An equipment list is presented in Table 8.143; a factored capital cost
summary is presented in Table 8.144; operating cost is presented in Table
8.145. The solution oxidizing potential certainly would be great enough to
Insure thermodynamic oxidation. The kinetics of such a rection, of course, are
unknown.
142
-------
TABLE 6.30 PROCESS COST: FIRST ORDER ESTIMATE
Unit Operation
1.
2.
Loach, jarostte
precipitation
Jarosite storage
Factored Capital
Cost Estimate
430,800
390.500
cosr is)
Annualized Capital Operation Cost
Cost Per Year
119.500 223.500
108.200 25.400
Total Cost
Per Year
343.000
133.600
3. Copper solvent
_, extraction, electro-
£ winning 336.100 93.100 205.900 299.000
4. Zinc, residual Iron
solvent extraction.
zinc sulfale crystal-
lization 661.600 183,300 269.700 453,000
5. Chromium oxid..
chromic acid pro-
duction 1.818.200 503,600 407.700 911.300
6. Nickel recovery 231,600 64,200 230,000 294,200
TOTAL COST 3.868,800 1.071.900 1.362.200 2.434,100
See Section 8.15 for details.
-------
TABLE 6.31 PROCESS COST SWHARV: FIRST ORDER ESTIMATE
Unit Operation COST ($)
Factored Capital Operation Cost Total Cost Potential Value
Cost/Vr 0 I2X Per Yr , Per Vr of Producl(t/lb)
-------
UNIT OPCMATION WAS XING-IKON 8X|
Ul
00 «.4 CPL/N
AlA COO LIT./LIT./H
KKSIDCNCK TIMK 1.4 H
nQTARV
KILN1
NIO
Figure 6.8a. S02/02 oxidation applied to chroalun oxidation and nickel recovery.
-------
me proposed possible application to the present system Is concurrent
oxidation of chromium (+3) and nickel (+2). The chromiun (+6) formed (Cr04")
would be present as an aqueous specie; the nickel (+3) hydroxide would be
present as a solid. Therefore, not only would the chromium be oxidized but a
separation between chromium and nickel would be achieved. In the present
flowsheet nickel and chromiun exist together. A treatment sequence could be
concurrent oxidation of chromiun (+3) and nickel (+2) at a pH ofNJ. Chromium
and nickel will be solid hydroxides at this pH. Therefore, the oxidation would
occur in a solid-solution slurry, Fhe research at INCO used calcium sulfite
(CaSO.) and oxygen as the oxidizing species. They proposed that the nickel
oxidation reaction was:
N1(OH), * CaSO, + 5/2 H.O + 3.4 0, >
'(solid) J{solid) c i
HI (OH), + CaSO "H-D
3(solid) * z
•
The oxidation was carried out in a modified flotation cell so that good
agitation and gas-solid-solution contact could be achieved. The measured
oxidation rate was approximately 5 g Ni**/!Her/hour at an equivalent S02 rate
of 7 g SQJliter/hour at a pH of 8; NI** concentration was 13.5 gpl, and an
oxygen supply rate of 600 liters/liter/hour was used. A similar chromium
oxidation reaction may be possible:
Cr(OH)3 + CaS03 + 3/2 H20 + 5/4 02 >
H£CrC4 * CaS04'2H20
Replacement of the electrochemical oxidation by SO.-O. oxidation and the
production of N10 rather than HIS results In a considerable potential cost
-savings. The process cost Is summarized In Table 6.32a. A comparison"Of costs
between the two flowsheets is presented In Table 6.33a. The ROI Is 411 for the
electrochemical oxidation flowsheet compared to 691 for the SO^-Og modified
flowsheet.
146
-------
TABLE 6.32«. PROCESS COST: FIRST ORDER ESTIMATE HODIFIEO FLOWSHEET HlCLUDING SO.-O.
tllROMIUH OXIDATION ' e
Unit Operation COST ())
Factored Capital Annual1 zed Capital Operation Cost Total Cost
Cost Estimate Cost Per Year Per Year
1. Leach, jaroslte
precipitation 430.800 119.500 223.000 343.000
2. Jarostte storage 390.SOO 108.200 25.400 133.600
3. Copper solvent
extraction, electro-
winning 336.100 93.100 205.900 299.000
4. Zinc,residual iron
solvent extraction,
zinc sulfate crystall-
ization 661,600 183.300 269,700 453.000
S. Chromium oxidation,
chronic acid
production, nickel
oxide production 1.043.900 289,200 484.600 773.800
TOTAL COST 2.862.900 793.300 1.209.100 2.002.400
-------
TABLE 6.J3a. COWARSION OF FIRST ORDER COST ESI I HATES SEIKEEN FIOMMIEETS FOR ELECTROCHENICM.
OXIDATION AND SO./Oj OXIDATION V CHROMIUM.
Flowsheet COST (})
F.C.C. F.C.A.C. Operating Total Product Value*
Cost/yr Cost/yr
Electrochemical 3,868.600 1.071,900 1. 367.200 2.434,100 5,641,400
(Table 2.1)
Hodlfied 2.862,900 793.300 1.209.100 2.002.400 :.88S,8CO
-------
Another alternative that appears to be attractive is presented in Section
8.15.7.4, solvent extraction of nickel by LIX63-D,EK?.' mixtures, e'ectrowinning
nickel, precipitation of chromium hydroxide, production of chromium oxide. The
anticipated unit operations a-e depicted in Figure 6.8b.
The solvent extraction of nickel from leach solutions containing chromium
appears to be possible by either use of a LIX63-D.EHPA organic or a D-EHPA-EHO
/44\ ' f.
organic* '. Preliminary shake tests were performed in this study. Tne
results Mere encouraging and verified literature data. Certainly further
research is needed to verify the conditions needed for an industrial SX system.
Also, one should be aware that solvent extraction of nickel using these
reagents is more risky than previously suggested alternatives because solvent
extraction of nickel (at low pH levels) is not yet practiced commercially.
The data on which the cost estimate for the modified flowsheet (Figure
6.7b) is made are presented in Tables 8.145 and 8.146. The process cost
summary is presented in Table 6.32b, and a comparison to the electrochemical
oxidation flowsheet is presented in Table 6.33b. The ROI is 41X for the
electrochemical oxidation flowsheet compared to 67% for the modified flowsheet.
Additional alternative unit operations are discussed in Section 8.15. The two
alternate unit operations presented in this section show good potential for an
excellent return on investment. Even if a credit is not taken for disposal the
modified flowsheets cost estimates show that the treatment process results in
income sufficient to offset the cost. It is recommended that further
consideration be given to these two flowsheets.
6.5. COMPUTER ASSISTED MASS BALANCE CALCULATIONS
6.5.1. Background
Rapidly escalating costs and constantly declining ore grades have prompted
the energy intensive metallurgical industry to seek new ways to improve process
economics. One of the methods that could be employed to immediately gain
greater operating efficiency could be the modernization of existing plants with
149
-------
NICKEL SOLVENT EXTRACTION
Froi Storage (Previous unit operation was
line-iron solvent eitraction)
pH
Lfl/Ul OR&AMLC-
Lf
f
••••••
**
a>
1
S
T
A
SIRIPPIO ORGANIC
t
H
S
— j
Lj
1
S
4
V
1
N
S
1
125 gal ll» 63
160 gal DlllPA
715 gal K(RNAC 510
660 gal NIXCR-SUIKRS
EXTRACTION
(1 CELLS)
RAfflUMt
pll . 1.0
v
NaOJ
PhCCNANT tlfCTROLTU
.STRIPPING —I
|t CELLS) I
NICKEL ELECTRO*INNINC
pn •<
IQ'.Oqal
pH
OEPKUD urciROim
NaOH (total: H> tpy)
PRCCIPMAIION VCSSUS
0.5 hr
Ni I ISO ppd
10 RCCVCU «CDS
CHROMIUM PRECIPITATION
IUM OXIDE PRODUCTION
pH . *
Figure 6.8h. Nickel solvent extraction, nickel recovery, chromium
oxide production.
ROIARY KUN
U90 ppd Cr 0
-------
Ul
TABLE 6.32b PROCESS COST: FIRST ORDER ESTIMATE MODIFIED FLOWSHEET INCLUDING NICKEL SOLVENT
EXTRACTION. NICKEL RECOVERY. AND CHROMIUM OXIDE PRODUCTION
Unit Operation
1.
2.
3.
4.
5.
F.C.C.
Leach, jarosite
precipitation 430.800
Jarosite ponding 3tO,500
Copper solvent
extraction, electro-
winning 336,100
Zinc, residual iron
solvent extraction.
zinc sulfate crystal-
lization 661.600
Nickel solvent
extraction, electro-
winning, chromium
oxide production 1,158.300
TOTAL COST 2.977.300
COST (t>
FCAC Operating Cost
Per Year
119.500 223.000
108.200 25.400
93,100 205.900
183.300 269.700
' 320.800 451.500
824.900 1.175.500
Total Cost
Per rear
343.000
133.600
299.000
453,000
772,300
2.000.900
See Section 8.15 for details.
-------
TABLE 6.33b COHPARilON OF FIRST ORDER COST ESTIMATES BETHEEK FLOWSHEETS tOR ELECTROCHEMICAL
OXIDATION AND NICKEL SOLVENT EXTRACTION AND RECOVERY.
Flowsheet FCC FCAC Operating Cost Total Cost Product Value*
Per Year Per Year
Electrochemical 3.863,800 1.071.900 1.362.200 2.434.100 S.643.400
Modified 2.977.300 824,900 1,175.500 2.000.900 5,977.100
ROI '1(5,977,100 • 2,000,900) / 2.977,300 1 (0.5)1100)
- 67 1 20 X
• Sane products In both flowsheets except for nickel (nickel In modified flowsheet) and
chromium (chromium oxide In Modified flowsheet).
See Section 8.15 for details.
-------
computer technology. Process modeling, microprocessor control and robotics
technology will play key and cost effective roles in process optimization.
Falling prices for all computer technology will enable even the smallest
company to benefit from these techniques. The main obstacle to this
computerization and optimization will be the availability of software specific
to the needs of the metallurgical industry.
Process modeling (especially mass and energy balance modeling) has long
been recognized as an engineering technique that enables metallurgical staff
members to design and operate efficient systems. However, these techniques
Involve many tedious, repetitive and time consuming calculations, and are.
thus, very labor intensive. Operating plants, particularly small operations,
often cannot afford to Involve engineers in such modeling even If plant
materials and energy are wasted In the process. Process modeling and/or
optimization could be a viabie technique for any operation if the lengthy,
repetitive calculations were coded into computer programs. Low cost, powerful
personal computers can make such process modeling an effective tool for each
engineer.
Mathematical process modeling of any metallurgical unit operation can
provide plant operators with an Incalculable amount of information concerning
plent practices. Mass balances can track the path of one or twenty or fifty
Items (such as metal ion concentrations) throughout the entire series of unit
operations. Recycle streams, changing flow velocities and mass additions can
turn simple mathematical calculations into a repetitious, time gobbling
nightmare. Keeping track of even one concentration or volume throughout the
entire series of unit operations can be extremely time consuming at best.
Changing one variable changes all calculations and starts the repetitious
process again. Tracking several Important values can be an Itemized accounting
mess. Process modeling of several interacting unit operations can
exponentially increase time consumption. It is simply too time consuming to
play "what if with the process model if the calculations are done by hand.
153
-------
Fortunately, the mathematical calculations involved in these mass and
energy balances are simply matters of repititious additions, subtractions,
multiplications and divisions. The process variables and the items to be
tracked can be often divided into a series of arrays. These conditions are
simply perfect for computer coding. Once the process calculation scheme is
developed, even i personal computer can trace several values at once. Disk
storage techniques can be employed to track an almost limitless (within reason)
amount of interacting items.
The research completed in this study investigated the use of an 8-bit
personal computer, an Apple II+, to model the mass balance calculations for the
extraction of metal values from mixed metal hydroxide electroplating sludges.
The models are, at this point in time, computerized mass balances that model
various extractive metallurgical unit operations. These models can be easily
adapted for optimization studies at a later date.
Several metallurgical unit operations were utilized in the extraction of
the various metal values contained In the electroplating sludges. Therefore,
several models were necessary to describe the research system. Also, these
models must "interact" so that the entire system could be researched. In othef
words, the outflow of one unit operation model would be the inflow of the next
unit operation model. The models were, thus, designed with this inflow/outflow
concept. However, es will be demonstrated later, this inflow/outflow concept
is an option to the computer operator. The operator can choose to have the
outflow of the last unit operation be the inflow of any of the listed unit
operations or provide a new inflow. This allows the user complete flexibility
within the complete series of unit operation models and makes "what if"
designing very easy.
This research completed the following computer assisted mass balance
models. It should be noted that these models were designed to describe a
specific extraction system and were not intended to be general models for any
system. However, modifications can be made to these programs fairly easily.
anu, they could be changed to define other systems as well. These models are:
154
-------
'Composite Sludge
This program allows the user to mix as many as 12 sludges together
to provide a composite sludge tnat will serve as the input sludge to
the leaching operation.
•Recycle Solids
This program allows the user to add recycle solids to the leach
vessel.
•Leach
This program models the leaching of the combined electroplating
sludges with sulfurlc acid and water. Three recycle streams may be
added to the vessel.
'Solid/Liquid Separation
This program models solid/liquid separations that involve filtering
and additions of wash water. Three different washing operations are
permitted.
'Solvent Extraction
This program models solvent extraction unit operations with a
maximum of three stages. The operator also has the option of
stripping the loaded organic.
'Precipitation
This is a general extraction model that allows the operator to
remove metal values from solution. T*»e operator may choose
precipitation of a species or may remove metal values with a "black
box" method so that the resulting stream may be the input to the
next operation. This is an especially useful method for "what if"
calculations.
All of the models monitor Important parameters with respect to 12 metals
Cu, Ni, Cd, Zn, Cr. Ca. Na. Fe. Al. Pb, Si and P.
6.5.2. Instructions
The diskettes and the instructions are provided as a separate document.
Example output of the calculations! program is presented in Table 8.147 for the
50 ton per day cost analysis flowsheet.
155
-------
SECTION 7
REFERENCES AND BIBLIOGRAPHY
7.1. REFERENCES
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•
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•
29. Fisher, J.F.C. and C.U. Notebaart. Commercial Processes For
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•
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«
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54. Campbell, M.E. and W. Glenn. Profit From Pollution Prevention.
Toronto, Ontario, Canada, 1982, 400 p.
55. Bonney, C.F., G.A. Gillett, and D.J. Everett. The Davy McKee
Combined Mixer-Settler. Its Commercial Performance. In;
Hydrometallurgy- Research, Development and Plant Practice, eds.
K. Osseo-Asare and J.D. Miller, Conference Proceedings, TMS.
AIME, New York, March 6-10. 1983. Atlanta. Georgia, pp. 407-418.
56. Boldt, J.R. and P. Queneau. The Winning of Nickel. D. Van
Noserand Co., New York, pp 487.
57. Verret, G. Personal Communications Oct. 19, 1984. Environmental
Resources Management, Inc.
162
-------
7.2. BIBLIOGRAPHY
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MITRE Technical Report MTR-7449, Series 3, Feb. 1977, 158 p.
Barthel, G. Solvent Extraction Recovery of Copper From Mine and
Smelter Wastes. J.M., July 1978, pp. 7-12.
Brooks, P.T., G.M. Potter, and D.A. Martin. Chemical Reclaiming of
Superalloy Scrap. USBM R.I. 7316, 1969, 28 p.
Oharwadkar, A.R. and N.D. Sy'. -ester. Factors Influencing Effectiveness
of Ion-Exchange Resins Used For Chromate Recovery. Ind. Eng. Chem.
Prod. Res. Oiv., 18(2):101-104, 1979.
Oevuyst, E.A., V.A. Ettel, G.J. Borbely. New Method For Cyanide
Destruction In Gold Mill Effluents and Tailing Slurries. Preprint:
14th Annual Operators Conference of the Canadian Mineral Processors
Division. CIM, Jan. 19-21, 1982. 14 p.
Oevuyst, E.A., B.R. Cr.nrad, V.A. Ettel. Pilot Plant Operation of The
INCO SO./Air Cyanide Removal Procass. Preprint: 29th Ontario
Industrial Waste Conference, Toronto, June 13-16, 1982, IS p.
Oevuyst, E.A., W. Hudson, B.R. Conrad. Commercial Scale Trails of The
INCO SO-/Air Cyanide Removal Process. Preprint: Canada/EC Seminar
Treatment of Complex Minerals, Ottawa, October 12-14, 1982, 15 p.
Devuyst, E.A., V.A. Ettel, G.J. Borbely and B.R. Conard. A New Proces
For The Treatment of Waster-waters Containing Cyanide and Related
Species. Preprint: 21st Annual Conference of Metellurgist, Toronto,
163
-------
August 29-September 1, 1982, 15 p.
•
Dutrizac, J.E. and 0. Dinardo. The Co-precipitate of Copper and Zinc
With Lead. Hydrometallurgy, 11:61-78. 1983.
Garrels, R.H. and C.L. Christ. Solutions, Minerals and Equilibria.
Harper and Row, New York. 1965. 450 p.
Jones, B.H. Recovery of Chromium From Tannery Waste. U.S. Patent
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JRB Associates, Inc. and CENTEC Corporation. Centralized Treatment of
Metal Finishing Wastes. Final Report EPA Contract No. 68-10-5052.
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Kershaw, M. and R. Pickering. The Jarosite Process-Phase Equilibria.
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T.J. O'Keefe, eds. Met. Soc. AIME, Feb. 24-28, 1980, Las Vegas, Nev.,
pp. 565-582.
Kopp, J.F. and G.D. McKee. Methods For Chemical Analysis of Water and
Wastes. EPA 600/4-70-020, March 1979. 430 p.
Kordosky, G., W.H. Champion, J. Dolegowski, S.M. Olafson, W.S. Jensen.
The Use of pH Control in Solvent Extraction Circuits. AIME Annual
Meeting, New Orleans, La., Feb. 19-23, 1979, 35 p.
Lee, C.K. and L.L. Tavlarides. Chemical Equilibrium Studies On The
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164
-------
Mansanti, J.G. The Precipitation of iron As A Jarosite From Iron, Copper,
and Zinc Containing Solutions. M.S. Thesis, Montana College of Mineral
Science and Technology, Butte, Montana, May 1978, 130 p.
Marino, M. Wastewater Control Treatment and Resource Recovery, hi:
Metal Finishing Guidebook and Directory Issue '83, J. Majio, ed.,
51st Issue, pp. 814-846.
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Sulfate-Sulphuric Acid-Water System. E/MJ, Oct. 1975, pp. 85-89.
Muller, M. and L. Witzke. Processing Nonferrous Metal Hydroxide Sludge
Wastes. U.S. Patent 4,151,257, Apr. 24, 1979, 5 p.
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Nilsen. D.N., R.E. Siemens, S.C. Rhoads. Solvent Extraction of Nickel
and Copper From Laterite-Ammoniacal Leach Liquors. U.S.B.M. R.I.
3005, 1982. 29 p.
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T.C. Lo, M.M.I. Baird and C. Hanson, eds. John Wiley and Sons, New
York, 1983, pp. 931-944.
165
-------
Preston, J. S. Solvent Extraction of Nickel and Cobalt by Mixtures of Carboxy-
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from Ammonical Solutions. Hydrometallurgy, 11:125-129, 1983.
Rice. N. M. Recommended Nomenclature for Solvent Extraction (Liquid-Liquid
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Solvent Extraction, Principles and Applications to Process Metallurgy.
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eds. John Mi ley and Sons, New York, 1983. pp. 945-954.
Tripler, A. B., Cherry, R. H., and G. R. Smithson. Reclamation of Metal Values
from Metal Finishing Waste Treatment Sludges. EPA 670/2-75/018, April 1975.
Valdes, C. J.. W. C. Cooper, and D. W. Bacon. Statistical Modeling of Solvent
Extraction Equilibria: Extraction of Copper from Sulfuric Acid and
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146:159-170, 1983.
Wikzke. L., and W. Muller. Process for Working up Nonferrous Metal Hydroxide
Sludge Waste. U.S. Patent 4.162,294, July 24, 1979, 6 p.
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pp. 919-930. " .
loo i
-------
SECTION 8
APPENDICES
8.1. ANALYTICAL PROCEDURES
8.1.1. Sludge Dissolution and Analyses
The dissolution of the sludge material was performed by the
following procedure:
*A 100 pound sample was removed from the barrel of sludge. The
sludge was cnooped up with a laboratory mixer Into pea size
particles.
"The sample was blended by repeated mixing.
'One hundred gram samples were split from the blended material
and placed In a drying oven at 100 C for 24 hours. Samples
were weighed and moisture content calculated. Samples were
run In quadruplicate to verify results.
'The dried material was ground In a mortar and pestlq to -100
mesh. A S.OO gram sample was split from the ground material.
'The dried ar.c sized sample was digested to determine
composition.
*A 5.00 gm sample was carefully weighed and placed in a
400 cc beaker. 100 cc of aqua regla was added to tf
-------
8.1.2. Aqueous Phase Analyses
Aqueous phase analyses were performed by ICP procedures. A solution
sample was withdrawn from a test sequence, filtered (if necessary), placed in a
25 cc polyethylene v
-------
Ot
IO
ABLE 8.).. PERFORMANCE EVALUATION REPORT:
Parameter
Aluminum
Cadmium
Chromium
Copper
Iran
Nickel
Zinc
12/03/82 MATER POLLUTION
STUDY NUMBER UP 009
Sample No.* Montana Tech True Value Acceptable
Foundation (mg/1) Limits
Valur(mg/l) (mg/1)
2
2
2
2
2
2
2
0
0
0
0
• o
0
0
.94
.OS
.31
.33
.14
.38
.45
0
0
0
0
0
0
a
.968
.072
.270
.338
.990
.400
.420
0.789
0.057
0.203
0.289
0.839
0.324
0.3/3
- 1
- 0
- 0
- 0
- 1
• 0
- 0
.220
.086
.330
.370
.110
.470
.462
Warning
Limits
(mg/1)
0.844
0.061
0.219
0.300
0.873
0.342
0.384
- 1.170
- 0.082
- 0.314
- 0.368
- 1.080
- 0.450
• 0.451
%
Performance
Evaluation
Acceptable
Check
Acceptable
Acceptable
Not Acceptable
i
Acceptable
Acceptable
* Calibration on Perkln Eljer 403 Atonic Absorption Spectrophotoneter set up for hundredth* of
mg/1 level; not calibrated for g/1. Therefore, all No.l samples not valid test of capabilities.
*• Iron hollow cathode tube defective.
-------
To a 125 ml Erlenmeyer flask add 50 ml H.O (deionized). 15 ml 20% H-SO^,
and 2 mis of H3?04. Pi pet in 1.0 ml of sample and add 8 drops of sodium
diphony1 amine sulfonate indicator. Titrate using moderate quantities of
dichromate while stirring or swirling. Proceed slowly with small volumes of
dichromate near the endpoint. When a violet color persists for one minute, the
endpoint has been reached. It is best to perform triplicate determinations
until volumes agree to within ^ .02 ml.
Report as gpl using:
P++ , (vol. O.C01N dichro-nate)(55.85)
e 9P TODS
The reagents required for Fe determination include:
•Potassium dichromate sdutlcn (0.2 N) made by drying pure K-Cr.O, @
120 C ana dissolving 4.9040 gm in one liter of deionized H.6. Dilute
to 0.001 N solution. '
"H2S04 solution; 20 v/o.
'Sodium diphenylamine sulphonate solution: 0.16 w/o.
•Concentrated H3P04.
The ferric iro_ij content of the aqueous phase was determined by
subtracting the Fe content from the total solution iron content.
Chloride a.id sulfate anion determinations were performed using a DICNEX
System 10 Anion Chromatograph. The procedure was: dilute the sample, 1/500,
with millipore treated deionized water; then analyze on the OIONEX system.
(Standards from 2000 g/ml S04* and Cl" should be made in such a manner to
bracket the concentrations in the sample).
Chromate or dichromate anion concentration was determined by the following
procedure:
'Determine the total chromium content of the solution by PA or ICP.
'Expose a known volume of solution to an equal volume r IRA 900 (a
strongly.basic ign exchange resin), "this quantitatively removes
the Cr04~. Cr^O^", or HCrO^" anions.
170
-------
Measure the volume of solution and analyze Che recovered solution
by AA or ICP.
'Vlash the resin with water and collect the wash solution, measure
the.volume and determine the total cnromium content of this sample
(Cr*3).
Calculate tne oxidized chromium in mg.
8.1.3. Organic Phase Analyses
Reagents:
Solvents: Two solvents are currently 1n use on a routine basis in
laboratories using ICAP Spectrometry on organic solutions; these are MIBK
(methylisobutyl ketone-spectral grade) and Xylene (mixed or para
Isomer-spectral grade). The first solvent i:, highly polar, meaning its
wettability of the sample uptake capillary is "well behaved" - similar to
water; however, It has a high vapor pressure and tends to produce an intense
emission in the plasma. It 1s also highly corrosive of most tubing materials.
Xylenes are non-polar and have a reasonably lower vapor pressure than rtlBK.
Since they do not wet the nebulizer in an acceptable fashion, the use of a
peristaltic pump 1s highly recommended to insure reproducible sample flow
rates.
Standards: Single element standards and a mixed element standard
containing 21 elements (S-21) are available from: Conoco, Inc., Ponea City,
Oklahoma. The standards carry the brand name conostan and can be purchased in
a variety of concentrations. It is highly recommended that single element
standards be purchased in order that spectral interferrons can be quantified
and stindard addition procedures may be employed in sample analyses where
necessary.
Procedure:
JY 48 Instrument parameters must ba set differently for aqueous and
organic solutions. The Incident power is increased and the sample flow rate 1s
reduced considerably (an order of magnitude for MIBK). Also, auxiliary plasma
gas is used to raise the bottom of the olasma one-half the distance from the
171
-------
Teflon support block (approximately equal with the top of the torch tip) to tne
first load coll. This prevents carbon buildup In the torch.
Instrument Parameters:
HI8K
Xyl enes
Incident
RF
1.5 Kw
4-5w Ref.
4-5w Ref .
Fine
Tune
7.1
3.5
Neb*
Pressure
20 psi
28 psi
Neb*
Flow Rate
.52
.50
Sample
Uptake
0.20 ml/min.
1.98 ml/min.
Nebulizer: Meinhart T230B2 concentric glass.
With typical solvent extraction type organic samples from metallurgical process
streams, the necessary dilution factor is 1/1000. This obviates any special
considerations one might have to give the sample because of unique physical
properties such as high viscosity, since the diluted sample Is mainly solvent.
Once the instrument parameters are optimized and inter-element corrections
quantified, analysis is as routine as aqueous samples.
8.2. SULFUR1C ACID LEACH STUDIES
8.2.1. Preliminary Testwork
Sulfuric add is a very effective lixlvant for treating mixed metal sludge
material. The design matrix and experimental results are presented
in Table 8.2. All experimental tests were run in a thermostated one liter
leach vessel under specified conditions of time, temperature, sulfuric acid
concentration. Eh, air purge, and agitation rate. One to two hundred gram
samples or undried sludge were leached with 250 cc of leach solution. Solution
samp'es were analyzed by Induction Coupled Plasma Spectrophotometry (ICP).
172
-------
TABLE 8.2. DESIGN HATRIX FOR SULFUR 1C ACID LEACHING OF SLUDGE (1/8 REPLICA): SERILS ONE
s
•
•
JIBl
JW
J9I
m
sin
V6
111
11?
•61
Batt
Unit
Hlyh (.)
lo. (•)
Icit lo.
?16 1
?
1
I
29? S
6
1
a
Bjielint
Effects
Cu
Fc
Cr
Hi
Zn
Cd
MM
(Hri.)
1.0
o.s
l.b
O.S
.
+
*
.
+
-
-
*
-0.2
1.1
0.7
-1.6
0
0
I»P.
<««
20°
...
BO"
20"
.
t
-
+
.
t
-
*
2.7
2.9
1.5
1.1
0
0
V°»
ISP))
90
20
110
70
.
.
*
*
.
-
+
*
0.5
1.3
0.7
-2.2
0
0
Ih
( X H«OJ)
5
5
10
0
.
.
.
.
t
t
t
»
2.0
1.9
1.5
2.0
0
0
Mr
Purft
No
«
Yes
No
_
»
*
.
.
t
f
-
0.2
0.9
0.7
0.6
0
0
>9il»U«B
RtU
(DPI)
370
..
SCO
?'°
.
t
.
*
»
-
t
-
0\2
u
1.5
2.7
0
Results: Extraction from Solid (X)
»4.. lip.
Cu
IU(M4
92.8
98.6
89.9
ittbai)
89.9
^5. 7
94.2
10.0'Ul
i6.0
Fe
BS.S (H.»
97 5
100
98.1
96.5 (96. S)
95.3
100
100
».«•:*. s
i6.0
Cr
9Lffiono)
ion
100
100
10*100)
97.0
100
100
II 1.^.6
iA 4
1 Nl
I6JJ90.J1
97.9
99.3
91.7
I?.V<87 i)
85.4
95.1
04.7
6«.?li.l
.•6 6
Zn
Lr.Xini})
100
100
100
100(91.1)
100
100
100
n. i:8.o
ilO 3
Cd
loonooL
Tofi
105 —
inn
nnlionl
'100
.00
ion
ii.i-i.i
•8 n
Volition
NOTE: -Sludge 5. Solids Coupes It Ion (t)
14. 52i0.25 Fe. 1.57i0.02 Cr.
3.17*0.03 Cu. 6.62i0.04 Nl.
9.68i0.08 Zn. 0.48i0.02 Cd
-Tests 2 and 5 duplicated
Baseline Run Three Times
-70 gpl Sulfuric Acid is Approxi-
mately the Sloicliionutric Acid
Requirement.
-------
The data in Table 8.2. illustrate excelle..; metal value extraction. The
effects portion of the table illustrates that the varia-ion In experimental
conditions chosen for study do not significantly influence metal value
recovery; e.g., for copper "the percent extraction is changed only by: -0.21
per 0.5 hr. increase in leach time; 2.7% per 60°C increase in temperature, etc.
The design matrix was repeated to consider the influence of acid content and of
sludge/liquid ratio. The results are presented in Table 3.3. The variable
(for the range studied) that shows the greatest influence on all metal value
extractions is acid content. Refer to the effects data in Table 8.3 for the
Influence of each variable on individual metal extractions.
The design matrix tests resulted in acid solutions that had pH values in
the range 0.5-1.5. A series of experiments were performed to investigate the
influence of pH on metal extraction. The results are presented in Table 8.4.
Two of the design table leach residues (Test 3 No. 291 and Test 6 No. 356)
were photographed and selected sections of the filtered solid were compared by
SEN analyses. The results are presented in F-gures 8.1, 8.2. and 8.3, Tables
8.5 and 8.6. Note in the photographs that the filtered solids contain a
variety of materials, e.g., sand-like particles, wood fibers, etc.
Leach residue samples were prepared for three of the design matrix tests.
I.e., samples were leached using the same conditions as specified in the design
table for Baseline (No. 261); Test 3 (No. 291); and Test 6 (No. 356)
conditions. Comparisons between the energy spectra for starting sludge and
leach residues are presented in Figures 8.4, 8.5, and 8.6. A quantitative
analysis of the same three residues is presented in Table 8.7.
•
The influence of time and solid/liquid ratio on metal value extraction is
Illustrated in Figure 8.7 and In Tables 8.8 and 8.9. Note that the leach
conditions for the test are baseline conditions and not optimum conditions.
However, the data do illustrate that the dissolution is very rapid.
A complete mass balance on a typical leach system was conducted to assure
that analytical results were reliable. The results are presented in Table
8.10. - 174.
-------
TABLE 0.3. DESIGN MATRIX FOR SULFURIC ACID LEACHING
s
•
1
m
Out,
UljJ
'bji
591
"sio
"BY
J3L
ill?-*
OF SLUDiiE (1/0 REPLICA): SERIES TWO
Bui
Unit
High (.)
I^IX
~i7u *
i
~"?"~
___")
t
5
6
;
8
1,-t
Effects (%r
Cu
Fe
Cr
Hi
Zn
Cd
lilt
(.m.)
3p_
"is
.«!>.-
:1§_..
r *
•»
-
4
-
»
2.2
8.5
5.1
-1.9
3.2
3.3
•IM»
(*of
Solidt)
55
25"
CO
...3
-------
TABLE
Semple
533
S34
535
536
537b
538
539
540
8.4. INFLUENCE OF Pll ON METAL EXTRACTION FROM ELECTROPLATING METAL HYDROXIDE SLUDGES
Condition
pH - 0.5
1.0
1.5
2.0
3.0
4.0
5.0
6.0
Cu
94.7
89.9
93.7
79.5
'49.3
13.5
1.7
0.5
Metal
Fe
97.4
91.3
92.0
46.7
0.6
<0.03
<0.03
<0.03
Extraction
Cr
99.4
94.7
96.5
71.8
17.0
5.1
5.1
5.4
from Solid (t)
HI
95.9
92.9
95.9
87.4
52.2
20.4
12.7
9.4
Zn
97.0
91.8
95.1
79.5
55 5 •
31.8
.4.1
2.2
Cd
93.0
93.0
93.0
84.8
69.7
46.5
23.3
<0.03
NOTE: -Sludge Barrel 1
Composition (I): 18.27*0.44 Fe, 7.84t0.40 Cu. 1.17:0.06 Cr.
5.53-0.33 N1.ll.4710.47 Zn, 0.73:0.04 Cd
•Solid content of sludge: 23.561
• 100 gm sludqe (*j;5J JJ js,°0ld$» slurrled In 200 cc H20 * X grams
•Time: 30 minutes
•Temperature: 25°C
-------
Figure 8.1 Photograph of residue from design matrix test no. 6.
177
-------
Figure 8.2. Photograph of residue from design matrix test no. 3.
178
-------
100X
Figure 8.3. SEH photomicrograph of section A from design matrix no. 6.
179
-------
8
TABLE 8.5.
Element
Cu
Fe
Cr
N1
Zn
Cd
SI
Al
Ca
S
P
Cl
SEH THIN SECTION ANALYSIS OF DESIGN MATRIX TEST THREE RESIDUE
ComDOSltlon (X)
Section B
3.59
8. 68
0.46
Vi
0.53
1.46
0
18.56
2.10
0.26
12.28
0.17
1.09
(TEST SAMPLE 291)
Section A
2.78
7.14
• 0.29
0.21
0.69
0
27.14
6.30
1.11
3.55
0
0.06
A: Leach residue 291 sample taken from the area marked A In Flgire 8.2.
B: Leach residue 291 sample taken from the area marked B In Figure 8.2.
-------
co
TABLE 3.6.
Element
Cu
Fe
Cr
N1
Zn
Cd
SI
A1
Cr
S
P
Cl
SEM THIN SECTION ANALYSIS OF DESIGN HATRIX
Comoosltlon (I)
Section B
3. SB
9.39
0.64
2.04
0.94
0
21.69
3.76
0.33
5.49
0.29
1.07
SIX RESIDUE (TEST SAHPLE 356)
Section A
2.18
7.79
0.29
0.72
0.3S
0
26.56
8.25
1.06
2.09
0
0.09
A: Leach residue 1356 sample taken from the area narked In Figure 8.1.
B: Leach residue 1356 sample taken from the area marked In Figure 8.1.
-------
5000-
COUNTS
CD
ENERGY IKEV) 10.0
Figure fl.4. Cornparsion of unleached barrel 5 sludge and leached sludge residue: •
Conditions (sample 356) given in table 8.2.
-------
5080-
COUNTS
CD
ENERGY 1KEV)
10.8
Figure 8.5. Comparslon of unleached barrel 5 sludge and leached sludge residue:
Conditions (sample 291) given 1n table 8.2.
-------
5388-
COUNTS
ENERGY (KEV)
18.9
Figure 8.6. Comparslon of unleached barrel 5 sludge and leached sludge residue:
Conditions (sample 261) given in table 8.2.
-------
TABLE 8.7. SLUDGE SOLID AND LEACH RESIDUE COMPOSITIONS
ffi
Oulgn
latrli
lilt*
ImllM «
(lilt fJIOl
ClOtfltlll 1
(lilt fill)
CtiXiltiii t
(liti fill)
lilttll
Ntgkt of
Solldi in
Sludgi
(<•)
41.44
41.44
41.44
riiii
Vilght of
Ruldui
((•)
l.tl
1.11
l.ll
Composition of Sludge Solids
fu
4.as
4.88
4.tt
r_
10. JJ
10.21
10. »
f-
1.19
1.19
1.19
HI
i.ia
i.ei
i.tt
i_
S.9{
S.«l
i.92
I.H
).)«
l.lt
Ml
7.11
Ml
4.i9
i.29
».«
f _
CA
i.n
i.u
i.n
o.u
O.S6
O.U
1.10
l.m
1.10
Composition of Residue Solldi
i.«
1.10
1.0
JM
t.»
Z.It
0.1S
0.11
0.10
0.5i
o.. t
O.SI
0.14
O.JO
0.0
0.16
.
o.u
o.u
Ml
I.SJ
10.11
>.!«
1.19
i.ei
l«
Cl
0.11
0.40
0.11
2.«>
I.Ct
2.10
0.17
0.40
0.1)
• See table B.2 for leach conditions. Residue 370 resulted from matrix test 1261; residue 371
resulted from matrix 1291; residue 372 resulted from matrix test OSS.
-------
TABLE 8.8. SLUDGE LEACI! TEST AS A FUNCTION OF TIME: BASELINE CONDITIONS TABLE 8.2
(C.6" SOLIDS)
Sample No.
357-1
357-2
357-3
357-4
358
(Repeat of 357)
358-1
358-2 *
358-3
358-4
358-5
Tine (nin.)
5
15
30
45
45
5
IS
30
45
60
Comparison to Design
Matrix Baseline
Conditions {Table 8.?)
Extraction (t)
Cu
71.0
75.2
79.4
77.6
79.6
77.2
78.3
77.1
75.9
78.8
70.0 i 4.8
Fe
65.5
71.9
76.2
75.7
76.9
73.4
75.9
75.1
74.3
77.1
74.6 ; 4.5
Cr
69.4
74.1
78.6
77.9
79.6
77.1
79.0
' 78.4
77.3
79.7
-81.1 t 3.6
Ni
64.4
67.8
71.6
71.1
72.3 '
70.3
72.4
71.6
70.5
72.3
69.2 t 5.4
Zn
71.2
76.7
81.6
60.7
81 .-9
78.7
80.7
79.5
78.7
81.6
77.7 t 8.0
Cd
eo.:
85.7
90.8
90.1
92.6
89.3
91.3
90.2
88.4
91.1
83.3 t 6.7
Notes: . 20°C. 90 gpl HjSO.
. 100 g sludge barrel 5/250 cc solution. 21.751 solids In sludge.
-------
TABLE 8.9. SLUDGE LEACH TEST AS A FUNCTION OF TIME: BASELINE CONDITIONS IN
TABLE 8.2 (16.21 SOLIDS)
Sample No.
252-1
2
3
4
S
6
7
8
9
10
Time (nln.)
10
30
SO
90
120
ISO
180
210
240
270
Cu
S4.3
G3.4
54.3
65.6
61.1
61.1
63.4
61.1
61.1
61.1
Fe
42.1
62.0
57.2
49.7
62.0
45.4
59.2
47.3
47.3
42.4
Extractlo
Cr
73.3
85. 6
65.2
73.3
4B.9
57.1
69.3
77.4
73.3
77.4
n (X)
HI
60.6
61.7
57.3
66.3
5B.4
71.6
66.1
65.1
71.7
58.4
Zn
50.3
45.8
45.8
4S.8
48.3
48.3
45.8
52.8
51.0
45.8
Cd
82.5
82.5
82.5
96.2
82.5
82.5
82.5
96.2
96.2
82.5
03
Notes: . 200C. 90 gpl MzSOa
. 200 g sludge. 22.72 S solids In sludge
-------
TABLE 8.10. MASS BALANCE ON LEACH *532
Conditions of leach: -1000 nm sludge in 1250 cc of leach solution (Barrel 2).
30 nin.
43-53°C
1302 of solids, i.e.. 163cc
•Time
•Temp.
•H2S04
•102 HN03
•Agitation:
1000 gm sludge (Composition: 18.27iO.445 Fe. 5.53:Q.33X Hi,
(23.195 solids) 2.80±0.1« Al . 11.47:0.47? Zn.
1.17± .061 Cr. 7.84±0.401 Cu.
0.73±0.045 Cd. 1.05±0.03S Ca.
4.54*0.445 P)
"I
28.97 nm residue of composition:
4.62*0.085 Fe. 1.69i0.035 Ni.
0.98:0.025 Al. 1.28=0.075 Zn. -S/L
0.26:0.02^ Cr. 1.26±0.085 Cu.
0.46:0.102 Ca. 0.4310.025 P
i- 178 qm HNO.J. 301.5 gm HZ
— 768 qm H,0
Solution Composition (diluted to 5 liters):
3.43 gpl Cu. 8.12 gpl Fe. 0.52 gpl Cr. 2.18 gpl Ni.
5.01 gpl Zn. 0.33 gpl Cd. 0.41 gpl Ca. 3.56 ',pl »
Element Weight
Material
Starting Solid
Leacn Solution
Leach Residue
Unaccounted (")
Cu
18
17
0
-1
.18
.15
.37
.6
Fe
42.37
40.60
1.34
-1.0
Cr
2.71
2.62
1.08
-0.4
Balance
NI
12
10
0
-10
.82
.90
.57
.5
( Grams 1
Zn
26
25
0
-4
.60
.05
.37
.4
Cd
1.69
1.64
0
-3.0
6
6
0
+4
Al
.49
.50
.28
.5
168
-------
B.6X Solids Initiilly
13.2X Solids Initially
Note in Table 8.2that baseline conditions
are not optimum for maximum recovery.
70
80
Time (Minutes)
Figure 8.7. Influence of Initial solid content on copper extraction from sludge as a function of
leach time: Baseline conditions table 8.2.
-------
Conditions for a standard leach test were chosen and all subsequent
testwork was based on these conditions: Temperature, 45-50°C, produced by
in-situ reaction heat and acid heat of dilution; time, 0.5 hr.; acid content,
130* of solid weight (this produces an acid solution in the pH range 0.5-1.5);
moderate agitation to suspend all solids in the solution phase; and a
sludge/liquid ratio of 0.8. A large number of leach tests, both in kettle
reactor and on a larger scale, confirm that sulfuric acid extractions are
excellent. A summary of a portion of these tests is presented in Table 8.11.
The leach procedure was found to produce pebble-like agglomerates of
unleached sludge if the sludge was slurried in water followed by addition of
acid then agitated. Extensive agitation failed to break up these agglomerates.
However, if the sludge was first exposed to concentrated acid then water added
•
to produce the desired acid concentration, agglomeration did not occur. The
leach procedure adopted consisted of blending the solids; adding the solids to
the reaction kettle; adding concentrated sulfuric acid (this raised the system
temperature to ahout 50-60°C) to the sludge; initiating agitation; adding
dilution water; then allowing reaction to proceed for one-half hour. All of
the sludge materials tested in this study responded well to sulfuric acid
leaching.
8.2.2. Large Scale Leach Testwork
Leach of 75-100 pounds of sludge in a single batch unit operation appears
to offer no chemical or mechanical problems. The extraction is rapid and
controllable. Excellent extractions are achieved for all metal values of
interest. Detailed experimental data for five large scale leach tests are
presented in Section 8.13. and are summarized in Table 8.11.
The test procedure is described in Section 5.1. Briefly it consisted of
blending a large sample of sludge material; sampling for moisture and chemical
composition determination; adding the sludge to a 120 liter or 270 liter
vessel; adding concentrated sulfuric acid slowly to the sludge; diluting with
tap water; and initiating agitation by an air driven one-horsepower agitator.
Reaction was considered complete after one-half hour. All of the large scale
190
-------
TABIE 8.11. EXAMPLES OF HETAL VALUE RECOVERY BY SULFURIC ACID DISSOLUTION
Sample No. Condition
533
534
535
942
532
2116
2621
2492
100
100
100
650
1.000
15.900
22.700
50.600
9.
9.
9.
9.
9.
9.
9.
9.
pH-0.5
pH-1.0
pH-l .5
pH=1.5
pll=1.5
pll-1.9
pH-1.5
pH-1.5
Fe
96.4
91.3
92.0
95. 4
95.8
62.3
65.0
92.0
Cu
94.7
89.9
93.7
94.9
94.3
75.9
92.0
93.7
Metal Extracted
Zn
96.9
91.8
95.9
90.5
94.2
83.8
96.9
95.1
HI
95.9
92.9
95.9
97.8
85.0
82.4
92.1
95.9
(I)
Cr
99.4
94.7
96.5
96.7
96.7
84.6
92.3
96.5
Cd
93.0
93.0
93.0
100.0
97.0
' 90.0
100.0
93.0
Al
89.
85.
87.
95.
96.
.90.
98.
96.
^^^B
9
7
1
7
0
3
6
9
Note: . All sludge samples were undrled.
. IbSOa added equivalent to 100X of solid weight
. One-half hour. 40-50°C
. Sludge/solution • 0.8
-------
leach tests were continued by changing the system conditions to precipitate
jarosite into the leach residue.
The results of the large scale test show that metal value extractions
achieved were very good and that a significant decrease in solids results,
i.e.. approximately an eighty-five percent decrease.
8.3. IRON REMOVAL (HIGH IRON SCARING SLUDGES)
Two major studies were conducted to investigate iron removal from leach
solutions containing high concentrations of iron (10-20 gpl) and low
concentrations of iron (<5 gpl); jarosite precipitation (8.3.1) and solvent
extraction of iron (8.3.2). The jarosite precipitation removal of iron is
conducted as the first unit operation after leaching and nay, in fact, oe best
performed concurrent with the leach process. The solvent extraction of iron
must be conducted after leaching, solid/liquid separation of the leach residue,
and solvent extraction of copper.
8.3.1. Iron Removal by Jarosite Precipitation
A commercial technique used for rejection of iron from a metal bearing
solution is the jarosite process (1,6,7). There are many forms of jarosite but
commercially either ammonium jarosite, .IH4Fej(S04)_(OH) • sodium jarosite
NaFe3(S04)2(OH)6; or potassium jarosite, KFe3(S04)2(OH)g, are produced. The
advantages of the jarosite precipitation process are:
1. Ferric iron can be removed from an acidic solution
(pH = 1.5-2.5).
2. The product is a readily filterable form.
3. The precipitation is selective toward iron.
Jarosites are a group of compounds having the general formula:
Ax(l-X)Fe3
-------
sodium or potassium ion is preferred in this project because of their
relatively low cost. The compounds most likely to precipitate from a sodium
containing solution are sideronatnte (Na2Fe(S04)2(OH)'3H20), metasideronatrite
(Na4Fe290°C); that the pH be maintained within the range 1.5-2.5; and that
residence times of several hours be allowed for good crystal growth (to ensure
good filtering properties).
8.3.1.1. Small Scale Kettle Testwork
A design matrix for sodium jarosite precipitation and the experimental
results are presented in Tables 8.12. and 8.13. As expected the Important
variables are solution temperature, starting pH, and residence time. The iron
in the-leach solution Is primarily ferric Ions so the Influence of hydrogen
peroxide is minimal. Note also (Table 8.12), that iron is the only element
appreciably Influenced by the changing variaoles.
A considerable number of jarosite tests have been performed. Illustrative
examples are presented in Tables 8.12. to 8.25. In general two approaches have
been taken: (a) the sludge was leached under standard conditions, filtered and
a jarosite precipitation study performed on the solutions or (b) the sludge was
leached under standard conditions, and a jarosite precipitation study was then
initiated "in-situ" leaving the leach residue solids in the reactor. The
advantage of approach (b) is that a product Is produced that is easily and
readily filterable; approach (a) requires that the leach solid be removed prior
to the jarosite precipitation. However, the leach residue is very difficult to
effectively filter.
The precipitation of iron as a sodium jarosite from a leach solution.
Table 8.14., can be compared to precipitation in the presence of the leach
193
-------
TABLE 8.12. DESIGN IVURIX FOR JAROSITE PRECIPITATION: EXPERIMENTAL DATA (1/4 REPLICA)
s
!
913
9U
9li
916
«»7.
91.6
919
sso
9i|-
But
Unit
High (.)
lo. (-!
Icil 1
\
2
3
t
SI S
6
?
a
Bui
Effect (gpl)
Fe
liip.
I°C)
90
— .
Boll in
80°
-
*
-
4
-
*
-
4
-1.6
list
(Hr».)
2
1
3
1
-
-
4
4
-
-
+
*
-0.8
"V*
dutch.)
2X
IX
3X
IX
-
-
-
-
4
4
4
4
0.2
Starting
pH
2.0
O.S
2.5
1.5
-
*
*
-
4
-
-
+
-1.1
"2°2
10 cc
10
20
0
-
-
4
*
4
4
-
-
0.0
Concentration In Solution (gpl)
•
Fe
8.95
S.lB
6.20
4. 83
7.68(8. SI)
7.15
9.15
2.35
7.SO(I.5S)
Cu
2.88
2.98
3.03
Z.88
;.9)(us]
2.08
?•'?
2.90
I.IB(».;O
Cr •
.28
•'3
1.20
i.oa
I.J5
1.25
.30
, -05.
1.30(1.28
N1
2.05
2.38
2.48
2.58
>.55C.S5
7.60
2.48
2.53
'.48(2. IS
Zn
5.70
5.65
5.83
5.80
S. '3(6. 00)
6.03
5.88
5.88
i.85(S.88)
Cd'
;
;
•
;
J5
,13
.28
.35
3.13(3.18)
3.43
3.28
_J
,33
3.73(1.20)
A1
2.00
1.75
2.18
1.93
2.00(2.281
2. JO.
2.23
1.85
2.20(2.2!
NOTE: -Starting solution composition (gpl)
•The leach solution was doped with Cr and Cd so
that their levels were significant and an
evaluation of possible co-precipitates could be
determined.
3.21 Cu, 9.C9 Fe. 1.27 Cr. 2.50 HI,
5.99 Zn, 3.34 Cd. 2.09 Al
200cc of Stirling solution, volume maintained
at ZOOcc by adding solution adjusted to desired
pH
Solids washed on filter with ZOOcc 5X
then diluted to SOOcc (data corrected to
original leach volume)
2 gro Jaroslte seed added
-------
(It
TABLE 8.13. DESIGN MATRIX FOR JAROSITE
^^
a
|
90
9tt
91 i
9(6
9*7-
9*8
9(9
9bO
ISI-2
PRECIPITATION: EXTRACTION FROM SOLUTION (1/4 REPLICA)
Bit!
Unit
High (.)
to. (-)
1(11 1
1
I
1
t
S3 S
6
I
B
Bin
Effect It)
he
lltp.
90
-
lolllm
BOU
-
+
-
*
-
«.
-
»
17.4
MM
(Hrt.)
2
1
3
1
-
•
+
+
•
.
»
»
8.9
(.Idled.)
2X
IX
3X
IX
-
-
-
-
+
«
*
t
-1.1
Stirling
pH
?.o
0.5
?.s
1.5
-
+
»
-
t
.
.
«
11.9
"A
lOcc
10
1?'
f
-
-
»
t
t
+
.
-
-0.4
Extraction from Solution (I).
Fi
43.1
32J
i
3
1)
1
47.0
16.0(6.0}
21.0
0
74.0
». 0(17.0)
NOTE: -Starting solution composition (qpl)
3.21 Cu. 9.09 Fe. 1.27 Cr. 2.50'N1,
c nn *— % <* * *»j A A A mm
vaw •••• " t »^ ««• f • • ««» ril
•200cc of starting solution, volume maintained
at 200cc by adding solution adjusted to
desired pH
-Solids washed on filter with 200cc 5S H2S04
then diluted to SOOcc (data corrected to
original leach volume)
•2 gm jarosltc seed added
-------
vO
Ot
Sample
696
897
398
903
901
TABLE 8.14.
Reaction
Time
(Mrs.)
0
1
2
2.5
2.6
REMOVAL OF IRON FROM SLUDGE LEACH SOLUTION BY SODIUM JAROSITE PRECIPITATION
Solution
nM
2.1
2.0
2.1
2.5
3.4
Concentration In Solution (gpl)
Fe
8.61
3.45
2.77
0.93
0.18
Cu
3.80
2.97
3.44
3.47
3.34
Zn
5.77
4.57
5.30
5.44
5.49
N1
2.69
2.14
2.52
2.59
2.66
Cr
0.58
0.38
0.42
0.29
0.10
Al
1.39
1.07
1.19
1.09
Oi47
Cd
0.39
0.30
0.36
0.37
0.38
NOTE: -Sludge Type A (Barrel 2)
•500 cc leach solution - exposed to Clj gas to raise solution Eh to 1164 MV.
•2.2 gm sodium jarostte added as seed
•Temperature: 86-92°C
•Raised pH el end of 2.5 hrs. to 3.4
•4 gm NaCOygm Fe
•Solution volume kept approximately constant
-------
residue, Table 8.15. The Initial rate of iron removal appears to be greatly
enhanced when performed In the presence of the leach residue. However, these
partlcrlar tests were performed at elevated solution Eh values and a large
fraction of the chromium was also precipitated, i.e., 83% and 86S. This effect
Mill be discussed unaer the sub-section (8.9) on oxidizing environments.
Copper and aluminum snows a decrease because of the formation of insoluble
•
phosphates (see Section 6.1).
Similar conclusions hold when considering larger scale testing, e.g.,
sodium jan>s1te precipitation from ten liters of solution showed 585 iron
removal in about 2.5 hours (Table 8.16) while in-situ precipitation showed 88i
(Table 8.17) iron removal in the same time period. Chromium loss to the
solids, however, showed a different result, i.e., for the precipitation from
solution only about 151 chromium loss occurred but in the in-situ precipitation
about one-half was lost to the solids (no external oxidizing reagent was
supplied to either of these precipitations). The chromium, copper, and
aluminum losses are a result of the formation of relatively Insoluble
phosphates.
• • i
Filterabillty of a leach residue product and a jarosite-leach residue
product 1s grossly different. I.e., filterability of a leach residue in a
filter press is extremely difficult (4.5 kg/m?/hr.) while filterabiHty of a
leach residue-jarosite product is much faster (22-55 kg/m /hr.) Qualitative
A
test results are presented In Table 8.18. Another factor that 1s important for
filterability is initial iron content. This effect is noted in the results of
a series of tests presented in Tables 8.19. and 8.20. In actuality the
formation of a poor filtering product at high Iron content Is most likely due
to the rate of addition of reagents rather than the Initial Iron level. The
conclusion of OutMzac* ' is that Jarosite can be effectively produced in
solution containing 0.25-3.0 M Fe* .
The choice of the precipitating alkali cation can be Na , K , or NH^ .
Precipitating testwork has been performed using each of the cations. Potassium
was chosen for the large scale test (but any of the three cations would be
appropriate) because according to Dutrizac' ' it forms the most stable
197
-------
Sample
899
902
904
Reaction
Time
(Hrs.)
1
2.5
2.6
TABLE 8.15. REMOVAL
Solution
P»
Fe
1.84 0.85
2.57 0.19
3.12 0.17
OF IRON
Cu
3.20
2.52
2.82
DURING IU-SITU
Concentration
Zn
6.44
5.27
6.01
LEACHING
1n Solution (gpl)
til
••••M^B
2.87
2.39
2.83
Cr
0.27
0.10
0.08
Al
1.91
I.Ob
0.69
Cd
0.40
0.33
0.38
NOTE: -Sludge Type A (Barrel 2)
•200 gm sludge 0 23.71 solids In 250 cc leach solution containing 23.7 gra J^SOa. Leached for
one-half hour at 8S°C before Jaroslte ppt was Initiated by adding
•4 gm NaCOj/gn Fe
•C12 added to raise Eh to 824 N.V. before NaCOj added
•Starting solution composition (gpl):
8.6 Fe. 3.8 Cu. 5.8 Zn. 2.7 Hi. 0.58 Cr. 1.4 Al. 0.4 Cd
-------
gpl
980
98)
982
983
984
98S
TABLE 8.16.
Condition
Starting solu-
tion pH adjusted
to 2.6
40 minutes
10S
140
I6S
Final
(t>H • 2.92)
JAROSITE PRECIPITATION FROM CONCENTRATED LEACH SOLUTION: TEN LITER TEST
Concentration in Solution
Fe
gpl I Ext
13.46
7.52
5.44
3.60
4.21
5.71 57.6
Cu
gpl I Ext
2.55
2.10
1.93
1.54
1.90
2.48 .2.7
Cr
gpl X Ext
1.38
1.05
0.98
0.73
0.87
1.17 15.2
(gpl) and Extraction from
HI
gpl I Ext
4.91
3.94
3.83
3.01
3.61
4.74 3.5
Zn
gpl I Ext
9.82
7.96
7.61
6.00
7.22
9.62 2.0
Solution (X)
Cd
gpl I Ext
0.48
0.37
0.36
0.28
0.34
0.45 §..£
Al
9Pl
5. 59
4.51
4.10
3.20
3.03
5.07
I Ext
9.3
NOTE: -Sludge type A (Barrel 5) leacl:«d 30 minutes under standard conditions at 82eC. Solids
removed and solution (10t) reheated to 94°C
•134 en N«C03 adJed and pH adjusled to 2.6
•Sample pulled as a function of time, volume of solution adjusted approximately to starting
level before sampling. Time samples, therefore, not considered precise.
•Volume adjusted during test period using water at pit • 2.6.
•25 gm jarosite seed added
-------
fs»
8
TABLE 8.17.
Sample
969
970
971
972
973
974
975
JAROSITE PRECIPITATION
Condition
Time
(mln.)
leach
40
Jaroslte
25
70
100
130
165
IBS
_J
1
EHj
.5
Fe
8.35
IH-SITU DURING LEACH PROCESS:
TEN LITER TEST
Concentration In Solution (gpl)
Cu
1.68
Cr
0.90
Nt
3.22
Zn
6.49
Cd
0.31
Al
3.
74
Precipitation
2
2
2
2
2
2
.6 Us-juiUd)
.7Udjuiti.il
.BUdjvlKd)
.7 (Wju.ledl
.7(«djmttd)
«
• •
3.07
2.48
1.06
1.31
1.37
0.98
0.91
1.52
1.56
1.85
2.55
2.76
2.19
2.12
0.65
0.55
0.49
0.66
0.71
0.51
0.49
3.15
3.01
3.75
5.46
5.69
4.40
4.28
6.31
6.19
7.70
11.00
11.53
9.07
8.80
0.31
0.28
0.36
0.54
0.52
0.42
0.41
2.
2.
'2.
3.
3.
3.
2.
93 .
88
74
64
07
07
92
NOTE: -Sludge type A (Barrel 5)
•5000 gm sludge leached 40 minutes under standard conditions before jaroslte reagents added
•10 liters solution, 137 gm Ha003, temperature 94°C
•Sample pulled as function of time, volume of solution adjusted approximately to starting level
before sampling. Time samples, therefore, not considered precise.
•Volume adjusted during test period using Mater at pH • 2.6
-------
ro
o
TABLE 8.18. POTASSIUM JAROSITE IN-SITU PRECIPITATION INTO LEACH RESIDUE:
RESULTS AND QUAl ITATIVE COMPARSION OF FILTERABILITY
Sample
1228
1230
1229
1231
1233B
1234
No. Condition
Starting Solution:
Standard one-half
hour leach
Four hour exposure
Series Repeat
Starting Solution
Four hour exposure
Comparative Leaches
One half hour leach
for filter companion
Repeat of 1233B
Fe
15.96
f rg*4iB'
1.74
[Fe**J-
14.34
2.01
[Fe**]-
14.81
[Fe«* -
12.16
[Fe** -
Concentration (gpl)
Cu Cr HI Zn Cd Al
5.46 1.00 2.83 8.45 0.39 6.23
1.3 gpl
4.36 0.6? 2.50 7.22 0.34 4.84
1.3 gpl
4.97 0.85 2.24 6.80 0.31 5.28
4.58 0.62 2.30 6.81 0.31 4.70
1.3 gpl
•
4.89 0.83 2.43 7.05 0.35 4.87
0.7 gpl
4.91 0.64 2.44 7.28 0.34 5.06
0.7 gpl
Dll FliterAbtllty
21 .....
1.7 Very Fast
2 2 '
1.7 Very Fast
1.0 Very Slow
1.0 Very Slow
Notes: . Sludge barrel 2 leached one half hour under standardi-condltlons(pll • 1.0) then jaraslte
conditions established.
. Filter comparsion conducted on a four Inch Mater aspirator filter. Qualitative filter rates are:
very fast - entire 900cc of leach solution cleaned of solids ln<5 minutes;
very slw - entire 900cc of leach solution cleaned of solids ln>l hour and several reptaccnents
of the filter paper required.
-------
FABLE 8.19. POTASSIUM JAROSITE IN SITU PRECIPITATION:
IRON CONTENT IN LEACH SOLUTION
Sample
2322
2325
2318
2321
2314
2317
Condition
15 gpl Fe Solution
Starting Solution
Final Solution, 6 hr.
10 gpl Fe Solution
Starting Solution
Final Solutloi, 6 hr.
5 gpl Fe Solution
Starting solution
Final Solution. 6 n.-.
Concentration
__C_u_
5.4S
4.24
3.SS
3.62
1.73
2.12
Fe^
14.96
~O5
10.31
TEH
5.60
uTT7
Zn
6.24
6.01
4.21
4.71
2.31
2.77
f
6.45
3.43
1.42
3.36
2.41
1.64
COMPARISON OF STARTING
(gpl)
N1
1.63
1.76
1.09
1.36
0.61
0.77
_Cd
0.08
0.08
0.04
0.05
0.02
0.02
Al
0.27
0.10
0.18
0.13
0.09
0.06
Filter ability
Poor
Excellent
Excellent
NOTE: -Each leach solution produced by varying the solid (Barrel 90°C, 6 hrs., 1 gm K^SOa/gra Fe, solution oxldiied with H20? (30*) during
last two hours of test, pH maintained 2.5 with KOH.
•The IS gpl test solution pH overshot to 3 3. Probable reason for high loss of copper In final
jarosited solution
-------
TABLE 8.20. POTASSIUM JAROSITE PRECIPITATION Of IRON AS A FUNCTION OF IRON CONCENTRATION
IN STARTING SOLUTION
Sample I Conditions
1056
1160
1161
1165
1166
1167
Starting Solution
4 Hr. Exposure
(unalterable)
Starting Solution
4 Hr. Exposure
Starting Solution
4 Hr. L'xp-JSurc
Fe
19.40
...
10.91
6.09
t
[Fe**] -
4.55
2.66
t
[Fe**] -
Concentration in Solution (gpl)
Cu
5.07
...
2.58
2.76
4.83 gpl
1.73
1.62
1.40 gpl
Ni
5.53
...
3.53
4.12
1.89
1.84
Cr
1.53
...
3.87
4.21
12.67
11.87
to
15.34
...
7.39
8.26
3.03
2.90
Cd
0.71
...
0.32
0.37
0.13
0.12
Al
8.22
...
5.05
5.41
2.05
1.91
NOTES: 'Leach solutions from standard leach on barrel 2 material; 1166 sample doped with high chromium leach
solution f'ooi barrel 8 material.
•750 cc solution adjusted to pH "2.0.
•21 gm seed.
• Temperature: 90"C.
•Final soli'!s fashed on filter pad. Extensive Hashing not performed.
•One gm KgSOj/gm Fe.
-------
jarosite. The present work indicates a faster rate of iron removal using
potassium. Tables 8.21. and 8.22. Additional jarosite precipitation data
summary tabulations are presented in Tables 8.23.-8.25., i.e.,
'Potassium jarosite In Situ Precipitation into Leach Residue: (Table 8.
23).
'In-situ potassium jarosite precipitation at initial pH = 3 (Table 8.
24).
'Comparative iron oxidation and jarosite in-situ precipitation (Table 8.
25).
8.3.1.2. Large Scale Tettwork
Solution iron content can be effectively lowered to the range of several
hundred parts per million by potassium jarosite precipitation. It would be
desirable to have a lower value but lower concentrations do not appear to be
achievable in large scale testwork. The presence of several hundred parts per
million iron Is not a major problem because it will be coextracted with zinc
during solvent extraction; will not contaminate the zinc strip solution because
it will not be stripped by 200 gpl H2S^4; and can be str'PPed from an organic
bleed stream to rejuvenate the organic.
Detailed results for five large scale sequential tests are presented in
Section 8.14. A summary of metal value loss from leach solutions are presented
i'n Table 8.26; 5-7 hours of precipitation time resulted in 94.4-99.5% iron
removal; 10.6-13.3% copper loss. 0-5.61 nickel loss. 25.0-42.8% chromium loss,
0-2.5% zinc loss, and 1.9-29.3% aluminum loss. A portion of the metal value
loss is recoverable by acid Teaching. This was demonstrated by leach tests
performed on the jarosite product from sequential test series five. The
composition of the jarosite residue is presented in Table 8.27. and the results
of leach tests on the residue are presented in Table 8.28. Three fourths of
the copper is recoverable, two-thirds of the zinc, 18.8% of the chromium and
100% of the nickel by an acid leach (pH =» 0.5).
204
-------
IM
s
TABLE 8.21. JAROSITE PRECIPITATION OF IRON FROM
Sample 1 Condition
1177
1181
1182
1183
1189
1190
1187
1188
NOTE: •
Starting Solution
NHd*, 4 Hours
Ha*. 4 Hours
K+. 4 Hours
Starting Solution
NHd*. 4 Hours
Starting Solution
K*. 4 Hours
Leach solution from barrel
Fe
11.53
5.28
9.80
2.39
10.90
5.23
t
[Fe*»] -
11.22
3.01
t
[Fe**l •
2 material
Cu
3.09
3.15
2.82
2.70
3.40
3.35
0.09 gpl
3.27
3.30
0.07 gpl
A NOMINAL TEN GPL IRON SOLUTION (OXIDIZED)
Concentration
HI
3.95
4.18
3.61
3.61
4.72
4.79
4.50
4.93
Cr
3.05
2.95
2.81
2.41
3.28
3.07
3.21
3.01
(qpl)
Zn
7.26
7.67
6.55
6.70
7.94
7.93
7.70
8.27
Cd
0.30
0.30
0.26
0.28
0.39
0.40
0.37
0.42
Al
4.64
4.63
4.13
4.08
4.85
4.65
4.78
4.76
-750cc of solution. pH adjusted Initially to 2.0.
t
•
•
"
21 g seed test.
1 gm reagent/ 1 gm Fe.
Temperature - 90°C.
50 cc of 301 H202 added to
each test
sample.
"
-------
ro
o
01
TABLE 0.22. JAROSITE PRECIPITATION OF IRON AS A FUNCTION OF TIME: HH4* .
Sample f
1077
1108
1110
1112
1114
1109
1111
1113
1115
Conditions
Starting Solution
NHa* Test
1 Hr.
2 Hrs.
3 Hrs.
4 Hrs. (final after
filterlnq)
+
K Test
1 Hr.
2 Hrs.
3 Hrs.
4 Mrs. (final after
filtering)
K*
Concentration in Solution (gpl)
Fe
25.00
17.14
13.98
11.27
12.04
r +# .
[Fe* ] •
14.19
10.93
9.10
8.26
, V*
[fe ] '
Cu
8.67
8.28
7.90
6.91
8.43
7.7 qpl
8.39
8.18
7.93
8.07
7.7 gpl
HI
7.42
6.91
6.66
5.84
7.15
7.31
7.18
7.00
7.09
Cr
2.86
2.66
2.48
2.16
2.61
2.63
2.56
2.44
2.44
Zn
16.63
16.97
16.12
14.08
17.07
17.43
17.05
16.52
16.81
Cd
1.04
0.99
0.95
0.83
1.03
1.04
1.02
0.99
1.01
A1
6.93
6.94
5.54
5.65
6.C7
6.88
6.67
6.50
6.61
NOTES: -Leach solution from standard leach on barrel 2 material.
•750 cc. pll • 2 (initial). +
•Approximately 50 gm. seed, source of seed was from 1092 (for NH4 ). 1093 (K )
•Temperature: 05-92"C.
•Final filtering Included washing with pll 2 water. This was not extensive Mashing.
only enough added to recover solution to initial 750 cc value.
-------
TfBLC 8.23. POTASSIUM JAROSITE IN-SITU PRECIPITATION INTO LEAD! RESIDUE: INCREMENTAL INCREASE IN PH.
Sample Condition
Concentration (gpl) pH
Fe Cu Cr Hi Z»L _Cd___ Al
1258 Starting Solution:
Standard 1/2 hr.
leach
11.51 3.33 0.62 1.53 4.95 0.21 4.52
lFeM) • 0.5 gpl
1259 2 hr. exposure 3.03 3.12 0.53 1.54 4.98 0.21 4.43 Initial • 1.52
Initial pH • 1.5 , Final - 1.28
[FeT*] - 0.05 gpl
ro
3 1260 2 additional hours ex- 0.70 3.53 0.54 1.79 5.74 0.25 4.77
posure. initial pH • 2.0
[Fe**] • 0.1 gpl
1261 2 additional hours ex- 0.33 4.57 0.58 2.40 7.77 0.33 4.77 Initial • 3.25
posure, initial pH « 3.25 .. Final - 2.66
[Fe**] - 0.2 gpl
: -Barrel 2 sludge leached under standard conditions for one-half hour then conditions changed to favor
Jaroslte precipitation; I.e.. T • 88-92°C. 1 gn K^/go Fe. 25cc of 3W H202 added slowly.
-------
TABLE 8.24. IN-SITU POTASSIUM JAROS1TE PRECIPITATION AT AN INITIAL PH - 3
Sample Condition Concentration Igpl) Fllterabillty
Fe Cu Cr Ni Zn Cd A)
1278 Starting Leach (1/2 hr.) 9.20 2.97 0.50 1.20 MB 0.18 3.92
pH • 2.0
1281 6 hr. exposure to 0.35 3.72 0.35 1.84 6.31 0.27 3.35 Excellent
Jarosite conditions
G pH • 3 (final pH • 2.5)
o NOTE: -Barrel 2 sludge leached under standard conditions for one-half hour, then conditions changed to form
Jarosite precipitation.
•25cc of 30X H.O. added slowly during test period.
•Temperature • 86-90°C.
•1 gm K2S04/gm Fe.
-------
TABLE 8.25. COMPARATIVE IRON OXIDATION AND JAROSITE 1N-SITU PRECIPITATION
I ™^^^^^^^—^^^—— ™•"•""™^—i^^» «_—^^—••_«KB«__^_IV_B«^—W«B~^«l V^_^_^^MB—
Sample No. Condi tlon Concentration {gpl)
Fe Cu Cr HI Zn Cd Al
Oxidation by Cr20
7
1317 Starting Solution 8.86 2.88 0.53 1.31 4.47 0.19 3.78
[Fe**] • 0.36 gpl
1318 Six hour exposure 0.14 2.54 0.61 1.31 4.48 0.13 2.49
[Fe*'] < D.L.
o
*° Oxidation by
1319 Starting Solution 9.57 3.25 0.56 1.32 4.54 0.20 4.07
[Fe**] • 0.28 gpl
1320 Six hour exposure 0.10 2.72 0.30 1.30 4.46 0.13 2.86
[Fe"] • 0.024 gpl
Notes: . Barrel 2 sludge leached one half hour under standard conditions, then conditions
established for potassium Jaroslte precipitation. Oxidant added dropwlse bcninnlng
four hours after start of test; 1.5 g K~Cr,0, added per gram of Iron. 25cc -30t H90,/950cc
solution. ' * ' ' z
-------
TABLE 8.26. METAL VALUE LOSS DURING LARGE SCALE JAROSITE PRECIPITATION
Sample No. Condition Metal Value Precipitation (%)
Fe Cu N1 Cr Zn Cd Al
-1371 Series One
Sequential Test
(Table 8.121) 99.5.13.3 0.0 25.0 0.8 0.0 1.9
5 Hr precipitation,
starting Fe; 14.00
gpl.
2125 Series Three
Sequential Test
(Table 8.125) 95.6 10.6 4.5 42.8 0.0 0.0 15.2
6 Hr precipitation,
starting Fe; 5.21 gpl.
2126 Series Four
Sequential Test (Table 8.126)
6 Hr precipitation, 96.3 12.7 5.6 41.0 2.5 0.0 21.3
after 8 hr settling.
2494 Serie Five
Sequential Test
(Table 8.127) 94.4 11.7 0.0 25.8 0.0 0.0 29.3
7 Hr precipitation,
starting Fe; 9.73 gpl.
Note: . Detailed results presented in Section 8.13.
-------
Sample No.
2611
2649
2650
ro 2651
265?
Average
Composition
TABLE 8.27.
JAROSITE RESIDUE FROM SEQUENTIAL TEST SERIES FIVE
Concentration In
Cu
2.74
2.80
2.73
2.88
2.88
2.81-O.C7
Fe
19.23
12.55
19.10
20.47
20.41
19.75-0.72
Zn
0.26
0.27
0.27
0.29
0.29
0.28-0.01
Cr
3.07
3.17
3.10
3.33
3.32
3.26-0.17
Solids (%)
N1
0.03
0.04
0.05
0.03
0.03
0.04-0.01
Ca
1.37
1.21
1.21
1.22
1.22
1.22*0.12
K
C.22
6.41
6.15
6.45
6.54
6.35-0.19
-------
TABLE 8.28. RELEACH OF PRECIPITATED JAROSITE RESIDUE
Sample No. Condition Extracted From Solid (%)
Cu Fe • Zn Cr N1 Al
2611 Starting Solid Composition *****
(Avg. Five samples) 2.81-0.07 19.75-0.72 0.28-0.01 3.26-0.17 0.04-0.01 1.50-0.06
2698 pH * 0.5 (Initial) 75.0 11.5 66.7 18.8 100.0 13.0
pl< « 0.7 (Final)
2701 pit * 1.5 (Initial) 25.0 S.9 ' 33.3 1.2 0.0 2.0
pH « 1.7 (Final)
2701 pH = 2.5 (Initial) 0.2 0.0 0.0 0.0 0.0 0.0
pll = 3.4 (Final)
Notes: . 10 g solid/100 cc solution, 25°C, 18 hour exposure.
. Sequential test series five jaroslte solid chosen for releacn study.
-------
8.3.2. Iron Removal by Solvent Extraction
It has often been stated that practically all metallurgical flowsheets
require a series of steps to successfully solve the problem of iron removal.
It Is true that In some mixed metal sludges iron is not an important .
constituent. If that is the case then the flowsheet presented in Figjre 6.1.
would be appropriate without the jarosite precipitation step. However, it Is
true that a plant that will be treating mixed metal sludge materials must have
a way to reject Iron. If the iron content Is high (a solution is produced that
contains several gpl Fe) then its segregation ran be via the jarosite (or
Geothite) process. However, the use of the jarosite precipitation process
results In an Iron bearing solution of a few hundred parts per million.
Removal of the residual Iron content Is required. Removal can be accomplished
by solvent extraction using one of two reagents, D-EHPA or Versatic Acid. A
major experimental Investigation was conducted on low iron bearing solutions
(that were not previously treated by jarosite precipitation). This study is
discussed in Section 8.4. The content of this section will, therefore, be
limited to Iron and zinc coextraction from jarosited solutions by O-EHPA. A
short discussion will be presented using Versatic Acid.
DgEHPA
Of the two reagents only DgEHPA is being used commercially to extract Iron
from a solution, e.g., Tecnlcas Reunldas uses a solvent extraction process to
(34)
remove the iron from a zinc leach liquor at its Espindesa operation*1 . Their
flowsheet is presented in Figures 8.8. and 8.9.
The anticipated selectivity of D-EHPA for iron is shown in a qualitative
way In Figures 8.10a and 8.10b. The figure suggests that Iron should be
selectively extracted from other metal values at a pH 1 or Iron and zinc
should be co-extracted from the other metal values at pH*>* 1.5.
O.EHPA was Investigated as an extractant for iron early in the present
experimental study. A design matrix Is presented in Table 8.29. The design
matrix verify, on a small laboratory scale, the selectivity of
D2EHPA-Oeconol-Kerosene mixtures for Iron. This conclusion is further
213
-------
RlflBUW -1-
Iratum to Inch tl
M*
MnfffiMt- -
.
Chlortfa liquor
from cMondma pynttt
enocrt leach
Sol««nt rvrr*cTion
i
1
Znd,l.np
J *
ZnO,
•QtMOV
MlullO
Zn nincnon
•nth 02EHPA
.
ZnnnppMig
with
L_
r
EkctrowuMMng
t
•
1B^«^«^B__
1 t
Ft «»lr«etion
OTIti R,NH,CI
1
n r«o,
n +
NO
Ft fnt 02EMPA
O2EHPA bind . MCI
Horgmc
ramovH
I
•
Zn
Figure 8.8. Outline flowsheet for the Espindesa process. (From Flctv35')
214
-------
. FeCljSOLUTION
*"" TO OKH LEACH
ro
ui
OEHiPA BACK _
TO EXTRACTOrr
DEHPA BLEED
IRON REMOVAL
HCI MAKE-UP
HCI
UC1
RECOVERY
IRON STRIPPING
-WATER
HCI SOLUTION AMBERLITE LA-2
BLEEP TO
'HCI V/ASii
(34)
Figure 8.9. Auxiliary facility for Iron removal from DgEHPA. (From Nogueria1 ')
-------
Figure B.lOa. Equilibrium distribution for metals in 30" D.EH?A-Shel'sol
System. Extraction from single metal.sulfate solutions
of 5 g/liter metal. (From Thorsen*"'}.
too
234567
40
20
(30),
Figure S.lOb. Influence of pH on zinc extraction by D^EHPA. (From Henkle )
216
-------
TABLE 8.29. DESIGN MATRIX FOR DgEHPA EXTRACTION OF IRON FROM SLUDGE LEACH SOLUTION (1/8 REP'.JCA)
s
in
ii«
1)0
in
33?
)))
m
126
127
e...
Unit
High (.)
1 OK I-)
Int lo.
1
I
1
*
S
0
'
B
Ban
Effects
Cu
Fe
Cr
Nl
Zn
Cd
DIHPA
(*)
20
5
25
15
-
+
-
+
.
+
+
0
«6.8
0
0
-0.7
0 1
lironni
XERMAC
470-B
...
470-B
450
-
-
+
+
.
•
t
0
-9.8
0
C
-1.2
0
Die ino 1
(*)
10
10
20
0
-
-
-
-
t
*
»
0
-8,0
0
0
-2.8
o
Oil
Mil
(•in.)
2
1
3
1
-
»
t
-
*
-
•»
n
+1KO
0
0
-0.9
n
lit
i»p.
(°c)
40
15
55
25
-
-
*
*
t
t
.
a
9.5
0
o
-1.2
o
Iticb
Solution
pH
1.70
0.20
1.90
1.50
-
*
-
+
*
'
.
0
5.1
n
o
-n T
o
Results:
Rn. dp.
Cu
0
0
0
0
0
0
0
0
0
<4.3Z
Fe
10.2
71.3
169 IV
33.3
45.'
42.1
0
10.0
22.fi
Viriiblt
f
Ektractton from Solution (S)
Cr
0
0
0
0
1
I
)
)
0
0
0
ib.l
. Nl
0
6
C
0
0
j
0
0
0
tS 0
7n
11.1
4.8
!.« ().))
3.5
0
0
0
0
2 0
i5 5
Cd
D
I)
0
0
0
0
0
0
n
it 0
NOTE: -Sludge Type A
•Initial Solution Ccnposttion (gpl):
0.04 Cu, 3.90 Fc. 0.49 Cr, 2.24 Nl ,
0.21 Cd
•Organic/ Aqueous • 1; 50 cc each
Test 5 Duplicated
.Observations on Phase Separation
Presented in Appendix Table 8.27.
-------
TABLE 8.30 OBSERVATIONS ON PHASE SEPARATION: DESIGN MATRIX TESTS (TABLE 8.29)
FOR IRON REMOVAL US I KG D^HPA
Test t Observations
I Some Muck*
2 Poor Separation, Very Kucky
3 Good Separation
4 Fair Separation
5 Poor Separation, Very Mucky
5b Some Muck
FN>
o» • 6 Good Separation
7
8
Baseline A
Baseline B
Baseline C "
•Muck • A Isyer of organic-aqueous that disappears slowly.
-------
substantiated by data presented in Figure 8.11 depicting the influence of pH on
iron extraction for a single contact shake test. A large number of extraction
shake tests were conducted and the results are reported in Section 8.3.3.
The extraction study was followed by a study of the stripping
characteristics of the organic phase. The strip tests immediately showed that
ferric iron 1s very effectively extracted from the leach solution using D-EHPA
but cannot be stripped from the organic phase by sulfuric acid. The detailed
results from an extensive series of tests are presented In Appendix 8.3.3.
The conclusion from the testwork is that iron cannot be stripped from
O.EHPA by sulfuric add. This result 1s in agreement with literature
sources but 1s not in agreement with patents by Reinhardt' .
Iron, however, can be effectively stripped from D.EHPA by hydrochloric
1341 '
add* . Experimental results are presented in Table 8.31. The ability to
strip Iron (and aluminum) from the organic pnase means active reaction sites
can be regenerated and the organic phase can be recycled for further iron and
zinc pick-up.
Zinc can be stripped in preference to Iron from a DgEHPA organic phase
using sulfuric acid. This result Is presented in Table 8.32., i.e., zinc Is
effectively stripped by sulfuric acid but iron is not stripped. The importance
of the fact that Iron and zinc bearing organic solutions can be selectively
stripped of its zinc content by use of sulfuric acid and then stripped of Its
Iron content by hydrochloric acid is that solvent extraction can be applied to
a mixed metal solution (Including iron). This fact allows an appropriate
treatment scheme to be developed for iron-chromium sludge materials without
using a jarosite precipitation unit operation: i.e., for a high chromium
bearing sludge the treatment sequence would be: leach; SX of copper from the
aqueous phase with an oxlme reagent; SX of iron and zinc from the aqueous phase
using DpEHPA; selective stripping of zinc from the organic phase by sulfuric
acid followed by stripping of iron from the organic phase by hydrochloric acid
solution.
219
-------
100
80
60
u
ID
O
40
20
1
25 v/o DF.HPA
10 v/o Iscdeconol ^j(
65 w/o Napcleup^*^
/
Contact Conditions:
PH . 1.5
K Eleaent Cone. (gpl)
Fe 0.95
Cu 0.026
Cr 0.^3
Ni 1.52
Zn 2.10
— Cd 0.16
1
__^^^
*~
t.t(i)
73.3
0
0
0
0
0
__ • '
pH -1.75
lone. (gpl) Cit(t)
0.26 92.5
0.027 0
0.47 0
1.68 0
2.18 0
0.17 0
•• i W
••
9* • 2.0
Cone. (gpl) EitU)—
0.062 98.2
0.028 0
0.09 80.8
0.40 75.6
0.55 /4.9
0.11 26.5-
Starting Solution Composition
™
1
3.55
0.028
0.47
1.64
2.19
0.15
1
0.22 gpl Fe
0.02 gpl Cu
0.05 gpl Cr
0.10 gpl Ni
0.14 qpl 2n
0.0? qpl Cd
1
0/A . 1
leap • 25°C
lice • 3 Bin.
1
1.50
1.75
Initial pH
2.00
Figure 8.11. Influence of pH on iron extraction by DjEHPA.
220
-------
TABLE 8.31. FERRIC IRON STRIPPING FROM DgEHPA WITH HCL
Sarnie Ho. Condition Concentration (gpl)
Fe Zn Cu_ Cr tH Cd Al
1322 Starting Solution 4.09 2.85 1.16 0.29 1.04 0.14 1.54
Iron Loading
1323 Contacted with OCHPA
(20 v/o); 0/A • 2,
rv> 40°. pH « 2.IB 0.32 1.75 1.15 0.28 1.04 0.14 1.37
fVJ
"* HC1 Stripping
1325A Contacted with
4N HC1; 0/A » 0.5, 40°C 0.80 0.26 0.004 0.002 < D.L < D.L 0.043
1325B Repeat 0.88 0.31 0.008 0.003 0.001 < D.L. 0.053
Strip Recovery (»)
1325A 8b_ 94_ — — — — 100
1325B J4_ lOQ — — — — 100
Notes: . All concentrations reported for the aqueous phases.
-------
ro
ro
Sample ho.
1466
1467
1468
1469
TABLE 8.32. BENCH SCALE
Condition
^
Jaroslted ( Barrel 14)
Leach Solution
Diluted 1466. pH > 1.75
Cu SX
LIX 622 (10 v/o) contacted
with 1467 (0/A • 1)
LIX 622 (10 v/o) contacted
with aqueous from 1468
(0/A • 1). Initial pH of
aqueous adjusted to 1.75
Zn SX: DEHPA (30 v/o)
Extraction
SEQUENTIAL SOLVENT
EXTRACTION TESTHORK: COPPER AND ZINC
Concentration
Ju Fe
3.14 1.44
1.56 0.69
0.02 0.69
0.004 0.68
Zn
^M^^PIMB
9.37
4.99
5.00
4.93
(gpl In Aqueous Phase)
Nl Cr Cd
4.95 0.54 0.52
2.75 0.27 0.24
2.74 0.27 0.25
2.71 0.26 0.25
REMOVAL
.-*'—
1.58
0.63
0.64
0.63
.
1470 Aqueous 1469 adjusted to 0.001 0.63 1.71 2.76 0.27 0.23 0.55
pH • 2; then contacted
(first contact) with DEHPA
organic (0/A * 1)
1475 Aqueous 1470 adjusted to
-------
TABLE 8.32. -CONTINUED
Sample Ho. Condition Concentration (gpi In Aqueous Phase)
Cu Fe Zn Hi Cr Cd Al
1477 Aqueous 1475 Adjusted to organic (0/A » 1)
to
OEHPA Organic Strip
1471 Organic 1470 (1st extrac- < 1 ppo <0.0) 3.50 <0.01 <0.01 0.02 0.01
tion contact) stripped
with 200 gpi H2S04
1473 Above organic (1471) < 1 ppm <0.01 0.01
-------
TABLE 6.32. CONTINUED
Sample No. Condition Concentration (gpl In Aqueous Phase)
Cu Fe Zn N1 Cr Cd Al
1476 Organic 1475 (2nd extrac- < 1 ppm 0.01 1.39 0.01
-------
Iron can also be stripped from D-EHPA by use of a reductive stripping
technique modeled after reduction stripping used in industrial uranium
recovery(37). The results are presented in Section 8.3.3. This technique
results in adding iron to the strip solution and, therefore, requires that a
disposal technique be developed for the strip (H-SO^) solution. The reader is
referred to Appendix 8.3.3.7. if interested in details of the study.
VERSA!1C ACID
Potentially iron can be removed by solvent extraction from an acid
solution by Versatic 911 - kerosene mixtures. An equilibrium distribution
diagram is presented in Figure 8.12. A selective separation of iron from Cu,
Cd. Ni, Zn appears possible using Versatic 911 as the extracting reagent. The
selectivity far iron requires the pH to be approximately 2.5. The problem with
this approach is that the iron product must be further treated, i.e., a
solution of ferric sulfate is produced by sulfuric acid stripping or a solution
of ferric chloride is produced by hydrochloric acid stripping. The advantage
of the previously discussed jarosite process over solvent extraction is that a
readily disposable solid product 1s produced in the jarosite precipitation.
Thorsen' ' and Teireira* ' has described a process for stripping Versatic
acid of its iron content by a process called hydrolytic stripping; see Figure
8.13. The advantage of the procedure 1s that solid Fe^O, is produced. The
disadvantage of the process Is that the stripping operation must be conducted
under autoclave conditions.
Experimental work was not performed in the present study using Versatic
Acid for iron removal nor is such a process presently used commercially.
8.3.3. Support Data; DpEHPA Load/Strip Testwork. Summary of
Testwork on_ Fe_ +_ 2n_ Extraction
8.3.3.1. Fe Extractions: Solution Preparation
The initial goal of this experimental work was to extract Iron from
copper-free raffinate using DjEHPA in 470B. A batch of leach solution (14 1)
was prepared for this study by the standard method, and then copper was
225
-------
100%
Figure 8.12. Equilibrium distribution for metals in 30% VERSATIC 911-
Shell sol system. Extraction from single metal sulfate
solutions of 5 g/liter metal. (From Thorsen(30)).
S'llphufic *•* «J N*'"t ftHttrOtyle)
Calcine (7x0)
Residue from
oiganic Icjchinj
Figure 8.13. Schematic flowsheet of integrated organic leach and
solvent extraction process in zinc hydrometallurgy.
(From Thorsen530'}.
226
-------
extracted in the full scale SX unit. Copper was extracted In two stages using
LIX-622 at 15 v/o. and stripped in one stage with 200 gpl acid. This technique
satisfactorily removed selectively the copper (Table 8.33).
The test provided operating experience with the large SX unit.
•
8.3.3.2. Fe Extractions: Influence of pH
Samples of the leach solution were adjusted to four pH values. 1.4, 1.6,
1.8, 2.0. They were then contacted for three minutes at an organic/aqueous
ratio of one using 251 02EHPA in KERMAC 4708 kero-ene at room temperature.
When carried out at this low temperature the extractions were quite good (as
indicated by pH change), but the phase separations were slow. By heating the
flasks to 38°C and re-mixing, the separations became rapid (less than two
minutes). The tests were qualitatively examined by determining the change in
pH that occurred in each contactor. These changes are indicated below. The pH
values at 38°C indicate that little additional loading occurred during the
second three minute shake period.
Original pH
1.4
1.6
1.8
2.0
Final pH at 25°C
1.28
1.32
1.35
1.41
Final pH at
1.28
1.32
1.35
1.41
38°C
The pH changes were converted to AH* and plotted in the graph below,
which indicates an abrupt increase in extraction at pH 1.8.
Experimental extraction data are presented in Table 8.34. Iron is very
effectively extracted at all pH values.
227
-------
TABLE 8.33. FIKST LARGE SYSTEM (I GALLON MIXER-SETTLER) TEST FOR COPPER
EXTRACTION USING LIX 622
r\»
IN)
OB
Conditions: 15 y/o LIX 622
85 v/o KERMAC 470B
Two Stages of Extraction
One Stage of Strip
pH of Leach Solution Into System: 1.75
Temperature : 25°C
Solution Flow Rate : 250 cc/mln.
Total Volume Treated : 14 liters
Strip Acid : 200 gpl HS
Cu
Original
Rafflnate
Feed (gpl)
(gpl)
2
0
.73
.043
Fe
6.
6.
10
14
1
1
N1
^^••^M
.90
.94
In
4.04
4.12
Jr_ Cd
0.42
0.42
0.
0.
24
25
-------
«o
TABLE fl.34. D,EHPA EXTRACTION FROM SLUDGE LEACH SOLUTIONS (COPPER
'FUNCTION OF PH
Sample pM
Start
589 (Starting
Solution)
585 1.40
586 1.60
£87 1.8-
588 2.00
Concentration In Solution
Final
1.28
1.32
1.35
1.41
Fe
6.18
0.18
0.11
0.08
0.08
Zn
4.34
3.73
3.67
3.39
3.68
Cr
0.41
0.37
0.38
0.36
0.39
FREE) AS A
(nnl) after Organic Contact
Nl
2.00
1.83
1.90.
1.78
1.97
^d^
0.27
0.25
0.26
0.24
0.26
Al
1.05
0.94
0.93
0.89
0.94
NOTE: -Organic: 25 v/o OEIIPA
75 v/o KERMAC 4708
•Contact Tine: 3 tnin.
•Temperature: 2S°C
•0/A - I. 50cc each
-------
.ozao
.0270 •
.02*0 -
.0290 -
.0240 •
.O210
Fe extraction series
Effect of initial pH
on [H+] transfer with
251 DEHPA in 4703
1.4 1.9 I •• I."
Initial pH
2.0
8.3.3.3. Fe Extraction: Influence of D.EHPA Content
Cu free raffinate (pH = 1.75) was contacted with an organic/aqueous ratio
of one at 40°C tfith 25. 30. 35. and 401 O^EHPA in 470B. Final raffinate pH
values were 1.41, 1.36, 1.34, and 1.35, indicating that 30% D-EHPA was probably
+++
sufficient to pick up most of the Fe . Table 8.35.
Two of the tests (251 and 40%) were repeated holding the pH constant
during the extraction by adjusting back to 1.75 after one minute of contact.
8.3.3.4. Fe Extraction: Stripping Series
The data obtained in the influence of pH test series and in earlier work
indicated that Fe could easily be extracted, perhaps not to 0 gpl in raffinate.
but at least to less than 1 gpl. It was assumed that 25% DgEHPA at 40°C,
230
-------
ro
TABLE 0 % INFLUENCE OF DEMPA CONCENTRATION OR LXTRACTION FROM SLUDGE LEACH SOLUTION
" ' (COPPER FREE) AT Pll - 1.75
Sample
589
590
S91
G9Z
593
594
595
UEHPA
(Starting Solution)
25 v/o
30 v/o
35 v/o
4u v/o
(Controlled pll)
25 v/o
40 v/o
Concentration In Solution After Organic Contact (gpl) Final
Fe
6.18
0.95
0.53
0.45
I
0.29
0.82
. 0.70
In
4.34
3.37
3.25
2.30
2.28
2.07
1.78
pH
Cr HI Cd Al
0.41 2.00 0.27 1.05 1.75
0.38 1.89 0.25 0.94 1.41
0.39 1.93 0.26 0.94 1.36
0.37 1.66 0.23 0.71 1.34
0.38 1.69 0.23 0.72 1.35
•
•
NOTE: -DEHPA aixed with Kenaac 4708
•0/A • 1, SOcc each
•Temperature: 40°C
•594, 595 controlled & pH • 1.75
-------
TABLE 8.36. SIMULATED CONTINUOUS LOAD/STRIP TEST USING 40X DEHPA ON COPPER FREE LEACH SOLUTION
OJ
ro
Sample
589
612
613
614
615
616
617
618
619
620
621
622
623
624
625
626
Description
Starting Solution
EH
E2
El3
I V
$11
$J2
S13
E21
E22
E23
521
S22
523
E31
E??
E3J
Concentration (gpl) after Organic
gpl In Aqueous
6.18
.55
1.53
4.41
0.42
0.24
0.30
~H(5 (81
4.77
4.78
5.29
0.73
O.S3
0.50
TT75 (12%
3.54
4.31
4.16
Fe
gpl loaded in Organic
5.63
4.65
1.77
1O5
stripped)
1.41
1.40
0.89
3T7J
of total stripped)
2.64
1.87
2.02 '
m
Contact (or stripping contact)
gpl In Aqueous
4.34
2.65
4.48
3.B4
1.53
0.06
0.03
Zn
Gpl loaded In Organic
....
1.69
-.14
0.50
7jB5
T76? (79.0S stripped)
2.72
3.92
4.74
1.34
0.20
C.04
T75B (761
3.90
4.05
3.80
1.62
•0.42
-.40
T3R
of total stripped)
0.44
0.29
0.54
TT77
•Temperature: 35-40°C
•pH (Initial): 1.75 for loading from aqueous phase •
•0/A • 1. 50 cc each
•Organic loaded via 3 contacts with fresh teach solution (E series)
•Organic stripped via 3 contacts with (A/0 • 1) 300 gpl ^$04 (S series)
•Three series of load and two series of strip tests conducted
-------
probably with some Deconol, would be suitable, and even though there was no
Isotherm available, three stages of extraction seemed reasonable.
Up to this time, however, there had not been many stripping tests carried
out, sc this series of tests was an attempt to understand stripping behavior.
A simulated continuous cycling test was carried out by contacting one
volume of 40% organic three times with Cu free 1.75 pH solution, the organic
phase was then stripped two times with 300 gpl acid. Extraction was repeated,
then the organic was stripped again, and finally loaded for the third time.
The results are presented in Table 8.36. pH changes in the rafflnates were
essentially similar for each of the three contacts and the analytical data
Indicated fairly good extractions. However, only a few percent of the iron
could be stripped.
8.3.3.5. Fe Extraction: Three-Stage Contacting
Copper free activated carbon treated (pH 1.7S) raffinate was contacted
three times with fresh 25 v/o (and a second test was made using 40 v/o) DgEHPA
organic. The pH change was measured on each raffinate.
The separations were quite rapid, and it was found that a small amount of
muck formed but if it was separated after the first contact, no further muck
developed.
Evidently most of the extractable iron (Fe ) was picked up in the first
contact, *ith very little more occurring in the second and third stages. Zn on
the other hand was progressively extracted; apparently Fe loads first, then
zinc is loaded. The results of these cests are presented in Table 8.37.
8.3.3.6. Fe Stripping
Each of the three organics from the 25% and 40% extractions were stripped
+++
with 40% H2S04 -(0/A * 10) in order to get some Fe color into the strip
solution. The phase separations were rapid and color did appear in the strip
•
233
-------
ro
TABLE 8.37. DEHPA EXTRACTION OF
Sample
i
657
651
652
653
654
655
656
Description
Starting Solution , pH=2.2
25 v/o DEHPA
1st Contact
2nd Contact
3rd Contact
40 v/o DEHPA
1st Contact
2nd Contact
3rd Contact
IRON AND ZINC: THREE STAGE CONTACTING
Concentration In Solution (gpl
Extraction from
gpl
5.64
0.86
0.64
0.66
Total
0.75
0.65
0.61
Total
Fe
% Ext.
84.8
25.6
0
Extracted 88.3"!
86.7
13.3
6.1
Extracted 89.29
) after Contact (0/A = 1
Solution (*)
Zn
gpl
3.85
3.13
1.87
1.27
Total Extracted
2.33
1.16
0.61
Total Extracted
1) and
% Ext.
46.5
40.2
32^0
67.0*
60.2
49.4
48.3
SOX
NOTE: -Copper free leach solution contacted three times with fresh organic.
•Temperature: 40°C. 0/A - 1, 50 cc each phase.
-------
from the first contact organic. Very littls Iron, but virtually all of the
zinc was stripped; Table 8.J8.
8.3.3.7. Fe Stripping: Variable Deconol Content
In previous work it had appeared that the presence of Deconol was
deleterious to Iron and zinc extraction but improved the stripping, so a series
of tests were performed with various contents of Oeconol in 251 D-EHPA.
First the organic phase containing 5, 10, 15, and 202 Oeconol was loaded
ntacting with Cu free
with 40% acid (0/A =• 10).
by contacting with Cu free rafflnate at 40°C. Then the organic was stripped
These data agree with the previous results that Fe stripping can be
Increased with Deconol but the degree of stripping is still poor; Table 8.39.
8.3.3.8. Reductive Stripping of Ferric Iron From O-EHPA
Removal of ferric Iron from O.EHPA can be accomplished by use of
hydrochloric add. It can also be accomplished by reductive stripping. A
number of experimental tests were performed to illustrate this fact. Tables
8.40 and 8.41 and 8.42.
If sulfurlc add Is contacted with iron powder; the iron is filtered from
the solution; then the solution 1s immediately contacted with the iron bearing
organic phase, the ferric iron Is stripped from the organic (Table 8.40). The
disadvantage of the process is that Iron is added to the strip solution.
Similar tests were concluded using zinc (Table 8.41) and nickel (Table 8.42)
rather than Iron. Neither was very effective.
A study was also conducted to determine 1f ferrous iron was extracted from
an aqueous solution by D.CHPA (Table 8.43). Ferrous iron was not extracted.
Strip tests using a reducing acid, i.e., sulfurous add, were performed on
ferric loaded O-EHPA. Removal of ferric iron was unsuccessful (Table 8.44).
235
-------
Sample
TABLE 8
.38. OEMPA STRIPPING OF IRON AND
ZINC (SEE TABLE 8.34 FOR LOADING DATA)
Concentration In Aqueous Strip (gpl) after One Contact (0/A • 10) and
Condition Extraction from Loaded Organic (%)
Fe Zn
i
658
G59
660
IN)
$
661
66?
663
25
Organic
Organic
Organic
40
Organic
Organic
Organic
gpl
v/o DEHPA
from 651 Stripped 0.29
from 652 Stripped 0.029
from 653 Stripped Sample Lost
v/o DEHPA
from 654 Stripped 0.14
from 655 Stripped 0.017
from 656 Stripped 0.012
? Stripped gpl SS tripped
1 6.45 89.6
1 9.23 73.3
Sample Lost
0.5 12.5 82.2
2 6.93 60.2
3 3.51 61.4
NOTE: '401 H2S04 strip used. Each of the six organlcs from Table 8.37 stripped.
•Temperature: 40°C
-------
TABLE 8.39. INFLUENCE OF DECONOL ON IRON STRIPPING FROM OEHPA
ro
to
Sample
SS7
€64
665
666
667
668
669
670
671
Condition
Starting Solution
5 v/o Deconol
Rafflnate
Strip
10 v/o Oeconol
Rafflnate
Strip
IS v/o Deconol
Raffinate
Strip
20 v/o Oeconol
Ruffincte
Strip
Concentration In Aqueous Strip
gpl
S.64
1.69
1.51
2.20
2.37
1.14
6.58
1.28
6.75
NOTE: -ATI tests used 25 v/o DEHPA. variable amounts of
•Temperature: 40°C
.Copper free leach solution contacted (0/A • 1);
401 H.SO. (0/A • 10)
(gpl) and Percent Stripped fron Organic
Fe
1 Stripped fron Organic
3.8
•
6.8
14.6
15.5
Oeconol. remainder Kermac 4/0 B
then organic stripped with
-------
ro
TABLE 8.40. REDUCTIVE STRIPPING OF DEIIPA. PRELIMINARY TESTS: IRON RE DUCT ANT
Sample No.
1336
1338
1339A
1339
Condition
Starting solution
oxidized Mlth H,C,,
pH • 2.22 e l
Contacted (1336) 20 v/o
DEHPA, 80 v/O KERHAC
4708. pH • 2.2, 0/A - 1.
2 minutes, aqueous phase
(pH • 1.04)
Loaded organic phase
200 gpl H2SO* contacted
with iron powder 5 minutes
filtered
1339A solution contacted
with Fe loaded OEHPA
Stripped (I) from OEHPA
Concentration (gpl)
Fe Cu Cr Hi Zn Cd Al
4.01 1.05 0.28 l.CO 2.74 0.12 1.40
0.85 C.96 0.25 0.92 2.09 0.11 1.24
3.16 0.09 0.03 0.08 0.6S 0.01 0.16
2.25 < D.L. 0.006 0.002 < D.L. < D.L. < D.L.
•
6.83 0.003 0.011 0.005 0.49) < D.L. 0.048
100 76
(complete)
Notes: . 50 cc of oxidized leach solution contacted with 50 cc of 20 v/o DEHPA. 0/A • 1, pll • 2.2
. Loaded organic (50 cc) contacted with 50 cc of 200 gpl H.SO. which previously been contacted
with iron powdwe. Contact with organic made iimedlately after iron-sulfurlc acid contact.
-------
IN)
U»
TABLE
Staple No.
1424
1415
1416
1417
8.41. REDUCTIVE STRIPPING
Condition
Starting Organic; 20 v/o
DEHPA. 80 v/o 470B;
Contracted solution 1322.
40°C. 0/A • 1, 2 rain.
Zinc Reactant
Solution 1424 contacted
with 200 gpl HjSOa (pre-
treated with 1 gpl In
then filtered: 0.16 SPl
Zn dissolved)
As above, except ml f uric
acid solution pre-treated
with 10 gpl Zn: 0.5? gpl
Zn dissolved
Strip Recovery (X)
Iron Reductant
Solution 142-! contacted
with 200 gpl H2S04 (pre-
treated with 1 gpl Fe.
OF DEHPA: SULFURIC
Fe Cu
OM ....
I1AIKI 1< ....
ACID COHI ACTED WITH ZINC AND IRON
Concentration (gpl)
Cr Nt 2n Cd Al
•
then filtered: 0.33 gpl
Fe dissolved
-------
TABLE 8.41. CONMNUED
Sample No.
lAta
I4IO
ma
Condition
• • jiKnuA Avf Ant ttt\fttftf
ns aoove except suiiunc
acid solution pretreated
with 10 opl re: 0.46 gpl
re dissolved
Ci> p 1 n DofnuAru 111
Sequence Test
Mrlnnnil Inxrfnil DrilPA
re Cu
1141)1 1« ....
(concentration (gpl)
Cr Ni Zn Cd
Al
Onno
.UU*
n me
with 200 gpl HiS04 to
remove zinc but not
iron
1420 Retreat above organic 0.07 O.'SS 0.008
Kith solution 1412
(10 gpl Zn pretreated
solution)
Strip Recovery (X)
31 ...» .... .... j7j .... ....
1421 Stripped loaded OEHPA 0.02 0.53 0.013
with 200 gpl H2S04 to
remove tine but not iron
-------
TABLE 8.41. CONTINUED
Sample No. Condition
1499 Qeftrajit Jihnu0 nrnjinlr
Concentration (gpl)
Fe Cu ^r
HI Zn
Cd
• »••
Al
n nin
«tth solution 1418
(10 gpl Fe pretreattd
solution)
Strip Recovery (S) 100% 4BX
NOTE: -All contacts 0/A • 1. R.T. 2 alnutes. SO cc of each solution.
fSJ
2 *Eh values for reductive solutions given In following table.
-------
ro
TABLE 8.42. REDUCTIVE STRIPPING OF OEHPA. PRELIMINARY TESTS: NICKEL RE DUCT ANT
Sample No.
1336
1338
1341A
1341
Condition
Fe
Starting Solution Oxldiied 4.01
with HjOypH • 2.22
Contacted (1336) with 0.85
20 v/o DEHPA, 80 v/o 470B.
pH • 2.2. 0/A • 1. 2
minutes, aqueous phase pH
• 1.04
Loaded Organic Phase 3.16
200 gpl HjSO* contacted 0.009
with nickel powder.
5 minutes, filtered
1341A solution contacted 0.77
with Fe loaded DEHPA
Stripped (t) 23.1
Concentration (gpl)
Cu Cr HI Zn Cd Al
I.OS 0.28 1.00 2.74 0.12 1.40
0.96 0.25 0.92 2.09 0.11 1.24
0.09 0.03 0.08 0.65 0.01 0.16
-------
TABLE 8.43. OEHPA EXTRACTION OF ZINC FROM A I*" -ft"
Sample
914
918
915
916
919
920
SOLUTION
Condition Concentration Remaining In Solution (gpl) and Extraction from Solution (I)
gpl
TotalFe
Starting Solution 12.00
Contacted Ml th DEIIPA 11.63
at pH • 2
Starting Solution 11.41
Adjusted to pH • 3.5
Contacted tilth OEHPA 10.78
at pH • 3.5
Starting Solution 10.80
Adjusted to pH > 4.0
Contacted with OEHPA 10.41
at pH • 4.0
NOTF: -75 v/o OEHPA. 10 v/o Oeconol.
•0/A • 2, 50 cc leach solution
•Temperature: 40°C
•Phase separation good
Fe*f
1 Ext.
Fe"
11.9
11.7 3.0
....
10.5 S.5
....
10.7 3.7
65 v/o Keraac 4708
. 100 cc DEHPA solution
Zn"
gpl X Ext.
9.00
4.99 47.5
8.66
3.86 55.4
8.19
3.83 53.2
-------
fABLC 8.44 . STUDY ON THE POTENTIAL SULFUROUS ACID STRIPPING OF IRON FROM OEIIPA
Sample
90S
906
907
908
909
910
9)1
91?
913
Condition
Loading Organic 0/A • 2
Starting Leach Solution
pH • 2
Contact 1, pH • 2.0
Contact 2. pH • 2.0
Stripping Organic A/0 • 2
10 v/o Sulfurous Strip
20 v/o Sulfurous Strip
40 v/o Sulfurous Strip
60 v/o Sulfurous Strip
80 v/o Sulfurous Strip
100 v/o Sulfurous Strip
Concentration in Aqueous Phase (gpl)
Fe
Aqueous Organic
8.8?
3.0/ 2.90
4.84 4.92
0.018
0.018
0.016
• 0.016
0.016
0.016
Zn
Aqueous Organic
5.97
4.91 O.SO
5.50 0.73
0.185
0.220
0.295
0.323
0.359
0.332
Cu Cr N1 Cd Al
3.95 0.60 2.84 0.41 1.41
4.0'. 0.61 2.88 0.41 1.40
3.92 0.60 2.85 0.41 1.38
•
NOTE: -Temperature: 40°C
•250 cc organic, 125 cc leach solution (Load)
•50 cc Sulfurous acid solution, 25 cc loaded organic (Strip)
-------
8.3.4. Iron Removal by Sludge Roasting
A few preliminary experiments were conducted to determine If iron could
be rejected from the sludge by an acid bake-roast-dilute acid leach (based on
(39)
Commercial L.R.E. process concept1 ')• i.e., the crncept Is to convert the
hyaroxides to sulfates in an acid bake, then to preferentially convert the iron
to an oxide form while leaving the other metal sulfates unoxidized. The
sulfate in the resulting roasted product would be water or dilute acid soluble
while iron would be insoluble. The results are presented in Tables 8.45-8.47.
Further testwork was deemed unnecessary.
8.4. IRON AND ZINC REMOVAL (LOU IRON BEARING SLUDGES)
The experimental work described in Section 8.3.2. led to the conclusion
that low Iron bearing solutions could be effectively treated without the need
for a jarosite precipitation unit operation. The major advantage of a
flowsheet tiat eliminatPS jarosite precipitation is that chromium Is not lost.
The differences In the low iron flowsheet and high iron flowsheet was
presented previously in Figures 6.1 and 6.3. The major difference in the low
Iron flowsheet is that jarosite precipitation of iron is eliminated and iron is
removed after solvent extraction of copper. Iron is removed by solvent
extraction using O.EHPA as the extractant. Iron is in fact removed separately
In a continuous solvent extraction system in conjunction with the removal of
zinc.
The experimental set-up Is presented schematically in Figure 8.14. It
consists of the following sequence of operations:
'Aqueous leach solution Is pH adjusted to the range 1-1.2, then
contacted with stripped organic (40 v/o O-EHPA, 60 v/o KERMAC 510
kerosene). The aqueous solution at this point In the flowsheet
contains Iron, zinc, nickel and chromium. Iron (2-4 gpl) is
extracted (to a few hundred ppm) from the aqueous phase into the
organic phase by one stage of mixing. Some zinc is coextracted but
chromium and nickel are unaffected.
'The iron loaded organic is transported to a second mixer where it 1s
contacted with 200 gpl H-SO.. Zinc is stripped from the organic
phase with the strip acia. Iron is unaffected.
245
-------
ro
IACLE 0. 45. PRELIMINARY DESIGN MATRIX AND EXPERIHENTAL RESULTS FOR ACID PUG-SULFAT10N ROAST (1/2 REPLICA)
5
a
?•
f
216
217
l\t
t\*
220
221
222
221
Effects
Cu
Fe
Cr
Hi
Zn
Cd
e»i
Unit
High (.)
U. (.)
lut 1
\
2
I
I,
•>
t
7
6
Kverag tt tl
58.2
4J.Z
33.6
45.1
52.5
75.8
Sulf*-
lion
ROM!
Ilit
'"riil
?
t
3
1
-
»
-
*
-
*
-
4
-3.2
-U.4
™ 1 • 1
0.4
-5.4
-2.8
Sulfitlor
"on!
lull
[»C1
600
200
800
400
-
.
•
«
-
-
»
4
-36.8
-42. 3
-3Z.6
-41.4
-10.7
-14.7
Solution
letch
(tip.
(•t)
RT
Boiling
RT
-
-
-
-
4
4
4
4
-3.0
U.B
1.0
2.5
-11.0
3.2
*cld
Content
of
Itieh
0
>t H?SOj
0
-
4
4
-
4
-
-
4
10.7
D.4
Ib.Z
6.4
-IZ.7
15.2
Results - Extraction from Solid (t)
Cu Fe Cr Hi Zn W
100.0 14.3 38.9 74. P 100.0 15.7
98.2 «4.9 91.4 92.5 79.? |8.?
45.6 0.2 0.0 :.7 55.7 67.3
0.9 0.6 0.0 0.4 19.2 2R.6
96.3 100.0 iod.O 100. ii 4.0 100,0
85.6 72.7 34.6 79.6 69.6 87.3
3.4 0.0 0.0 1.1 72.0 41.3
35.4 3.5 3.7 9.b 2C.6 87.:
HI., lip.
V (Hilton
NOTE: -Starting Sludge A (23.4 1 1.6X Solids)
•9.8 gn Cone. H2S04/150 gra Sludge; pug roasted
at 95°C for 3 firs.
•Sulfatlon Roast In 41 03. 81 SO;
•Solution Leach 30 Hin.. IOZ Solids
•Solution Diluted to 500 cc for Analysis
-------
TABLE 8.46. SULFATION AS A FUNCTION OF ROAST TEMPERATURE: EXTRACTION OF SOLID
•
Sample
957
958
959
960
961
962
Notes:
! Condition
Cu
No Roast 63.8
200-C 70.2
400aC 65.0
500°C 46.4
600°C 51.0
700°C 56.1
Extraction from Solid (I)
Fe Cr
75.8 85.5
72.3 65.0
61.0 56.4
•
37.0 27.3
29.6 17.1
7.7 5.1
HI
69.4
68.0
57.1
36.2
45.2
4S.2
. 10 a dry solid, barrel 2 sludge.
. Acid pugged with 10 cc H.SO. 0 95°C for 3 hrs.
. Baked sample roasted at Designated temperature In 41 0,. 8S
. Roasted sample leached in U H.SO. for 30 minutes, filtered
(data corrected to leach solution volume of 200 cc)
Zn
80.2
85.4
75.0
52.6
57.2
59.2
SO,. 88S
, washed,
Cd
68.5
68.5
60.3
41.1
54.8
65.8
N, for one hour.
diluted to 500 cc
Starting solid co.nposition: 18.29T, Fe. 5.531 Ni. 2.80S Al. 1.171 Cr.
0.731 Cd, 11.47* Zn. 7.841 Cu. l.OSX Ca
-------
TABLE 8.47. SULFATION AS A FUNCTION OF ROAST TEMPERATURE: DATA
ro
O
Sample
957
958
959
960
961
962
Condition
No Roast
200°C
«00°C
500°C
600°C
700°C
Concentration In 200 cc Solution (gpl)
Cu
2.50
2.75
2.55
1.82
2.02
2.20
Fe
6.92
6.60
5.58
3.38
2.70
0.70
Cr
0.50
0.38
0.33
0.16
0.10
0.03
N1
1.92
1.88
1.58
1.00
1.25
1.25
Zn
4.60
4.90
4.30
3.02
3.30
3.40
Cd
0.25
0.25
0.22
0.15
0.20
0.24
__A1.__
1.70
1.18
1.12
1.10
0.68
0.22
NOTE: >10 gn dry solid, Barrel 12
•Acid pugged with 10 cc H?S04 9 95°C for 3 hrs.
•Baked sample roasted at designated temperature tn 41 0», 81 SOj, 881 N? for one hour
•Roasted »ample leached In It f^SO^ for 30 minutes, filtered, Mashed, diluted to 500ml
(data corrected to leach solution volume of 200 cc
•Starting solid composition: 18.27X Fe, 5.53X Ni, 2.80S Al. 1.17S Cr.
0.73X Cd. 5.53X NI, 11.47X Zn. 7.84X Cu.
1.05X Ca
-------
r~
IS)
Figure 8.14. Zinc and Iron solvent extraction flow pattern: Phase II.
-------
'The iron loaded organic Is transported to a third, fourth and fifth
mixer wnere it is contacted witn 6 N HC1. Iron is stripped from the
organic phase into the strip acid.
'The iron, zinc depleted organic phase, is transported to a series of
three mixers where it is contacted with tne iron depleted aqueous
leach solution. The aqueous phase pH is raised to 2.0. Zinc is
extracted from the aqueous phase into the organic phase. Chromium
and nickel were unaffected. The raffinate is relatively free of
zinc (<50 ppm).
"The zinc loaded organic phase is transported to two mixers where it
is contacted with 200 gpl H.SO.. Zinc is stripped from the organic
phase into the strip acid. The organic is recycled back to the
first .nixer to contact fresh leech solution containing iron and
zinc.
8.4.1. Large Scale Iron and Zinc Removal
The results of a series of studies using the Reister one-gallon
r.ixer-settler solvent extraction rack are presented in Tables 8.48-8.49.
Earlier testwork snowed that a murk problem resulted in the first
extraction cell if the iron content was in the range of a few grams per liter
at a pH of about two. Follow-up testwork showed that the muck problem was
minimized by lowering the feed leach solution pH (decreased Iron and zinc
loading into the organic phase) and by changing the kerosene diluent (from
KERMAC 470 to 510, a low aromatic solvent).
v.
A series of coextraction tests were conducted using the solvent extraction
flow pattern depicted schematically in Figure 8.15. The results of two large
scale tests are presented in Table 8.48. The test results showed that
coextraction of iron and zinc was excellent and controllable and that muck
formation was minimized by running the firs: stage organic continuous instead
of aqueous continuous.
The above tests were followed by continuous large scale testwork over a
four day period. Sludge was leached, copper was extracted by LIX 622
extraction *nd the resulting solution was used for the four day test run. A
total of 365 liters of aqueous leach solution was exposed to 38.8 liters of
organic phase (Table 8.49). No degradation of the organic phase was noted.
250
-------
IM
Ul
TABLE
Sanple Mo.
3208
3216
3217
3255
3250
3251
8.48. LARGE SCALE IRON-ZINC EXTRACTION TESTWORK (PHASE II, LOU IRON PRELIMINARY TESTUORK)
Conditions Concentration In Aqueous, gpl
Fe Cu Zn Cr HI Al Ca Extraction
Efficiency. 1
Zn Fe
Large Scale Test on
co-extraction of Fc A Zn
where pH control exer-
cised In 1st and 2nd
cells (see notes below)
Starting solution. 75 1
1st Cell Feed. pH - 0.62 1.728 2.425 6.471 2.502 0.029 0.702
2nd Cell Feed. pH - 1.56 1.165 2.164 5.959 2.319 0.027 IO.B 32.6
4th Cell Final Raffinatc 0.011 0.048 6.153 2.829 0.014 — 3O 9970*
5 hrs. continuous test
Large Scale Test on
co-extraction of Fe ft Zn
where pH control exer-
cised in 1st and 2nd
Cells and the 1st Cell
was run organic continuous
First Cell Feed. pH - 0.95 1.611 — 2.231 5.470 2.547 0.035 0.696
Temp. - 22°C
1 Hr.
1st Cell Raf finale. 0.7475 1.935 5.664 2.602 0.036 0.652 13.3 53.6
pH - 0.83. 22°C
2nd Cell Feed. pH • 0.764 2.121 5.031 2.077 0.040 0.635
2.0). 2b»r
-------
ui
TABLE 8. 1C. tOMINUCD
Sample No
. Conditions
Concentration in Aqueous, gpl
Fc
Cu Zn
Cr
HI
Al
Ca Extraction
Efficiency. X
32b2
3256
3257
32b8
3261
3262
3263
4th Cell Raffinatc.
pH • 1.15. 21°C
3 Hrs.
1st Cell Raffina».e.
pll - O.U9. J7°C
2nd Cell Feed, pH -
1.91, 5U°C
4th Cell Raffinate.
pH - 1.34, 28°C
5 Hrs. (end of run)
1st Cell RofflnaU,
pll - 0.68. 30°
2nd Cell Feed, pll •
1.80. 58°
4th Cell Rof finale,
pH « '..19, 25 C
0.009
u.373
0.482
0.020
0.430
0.385
0.018
0.067
1.586
1.685
0.022
1.754
1.676
0.046
5.010
5.659
5.394
5.589
5.614
5.292
4.981
1.B70
2.587
2.448
2.392
2.558
2.430
2.149
0.005
0.032
0.026
6.005
0.034
0.031
0.012
Zn
0.220 96.8
0.692 28.9
0.661
0.322 98.6
0.722 21.4
0.707
0.317 97.2
Fe
98.8
76.8
95.9
73.3
95.3
NOTES: System set-up described In Section 8.4.1.
'Conditions of first large scale test: 75 liters of leach solution. 38.8 liters 401 DEMPA in 470-B, 18.2
liters of 200 gpl H.SO., 27 liters 4N HC1, flowrate - 0.25 l/m«n.. total exposure tine • 5 hrs.
'Conncnts on first large scale test: Some crud noticeable In Cell 12 1/2 hour after run began, but
formation remained minima) for this experlMcnt. Run demonstrated that crud formation can be
controlled by nil manipulation and still achieve satisfactory extraction of 3.5 g/1 combined Fe and
Zn. Iron Mas not stripped very we'l in this run.
'Conditions of second large scale te;t: 75 liters of leach sol'itlon, 38.8 liters 40 v/o DEIIPA In 470-B.
18.2 liters II.SO.. 27 liters 4.5U IIC1. flowrate • 25 litcrs/nln.. total exposure • 5 hrs.
•Conacnts on second large scale coextractlon test: organic continuous operation in 1st Cell mixer
exhibited positive results on minimizing crud formation for this node of mixer operation.
-------
ro
in
Barren Organic
SETTLER
IICI Strip
StobJ I
O
SETTLER
Extraction
Stage 4
MIXER
MIXER
Extraction
Stage J
SETTLER
SETTLER
Extractloi
Stage 2
Aquoous Feed
MIXER
Extraction
SI ago I
UETTLCR
O
Loaded Organic
Figure 8.15. Experimental set-up for Iron and zinc solvent extraction.
-------
Cell efficiencies for iron, zinc and iron plus zinc are reported in Tables
8.£0-8.52. Overall extraction efficiencies for end of run (EOR) solutions
showed: 97.4% iron extraction; 97.3% zinc extraction; and 95.1-99.2% iron plus
zinc extraction. Final solution raffinate contained: 0.053-0.095 gpT iron and
0.040-0.061 gpl zinc.
8.4.2. Continuous Long Term Solvent Extraction Testwork: Iron and Zinc
A series of solvent extraction studies were conducted to investigate:
Iron and zinc extraction stage and process efficiency; and possible degradation
of the organic extractant when exposed to a large volume of leach solution,
i.e., what is the effect on the organic extraction efficiency of a large number
of load/strip cycles. A schematic diagram depicting the flow patterns in the
test system for the aqueous and organic solutions is presented later in Figure
8.20.
The tests were conducted in the Bell Engineering testrack; 7.6 liters of
40 volume percent D.EHPA - 85 volume percent KERMAC 510 kerosene was contacted
with 150 liters of aqueous leach solution over a period of 67 hours.
Approximately 58 load/strip cycles were achieved. Over 232 loading contacts
and over 586 stripping contacts were made during the test period. An
aqueous/organic contact ratio of approximately 20 was achieved. The results of
the study are summarized in Table 8.53; stage efficiency and process rack
efficiency are summarized in Tables 8.55-8.57.
8.4.3. Crud Formation and Control During Iron-Zinc Solvent Extraction
8.4.3.1. Crud Formation in D.EHPA Solvent Extraction
A problem developed during the Phase II large scale iron and zinc solvent
extraction testwork. I.e., crud formed In the first cell of the extraction
(loading) stage and was initially uncontrollable. The flowsheet being tested
during that period was two stages of extraction at pH 2 followed by
readjusting the pH back to 2-2.5 after the second cell prior to entry Into
extraction cells 3 and 4. The formation of crud in the first cell created
several problems. Tne first cell interface was uncontrollable; crud was
254
-------
ra
in
ui
TABLE 8.49. LARGE SCALE IRON-ZINC EXTRACTION TES~UORK:
Saaple
3Z82-B
3282-8
3283-8
3284
3311
3312
3313
3314
3319
3320
3321
3326
3327
Conditions
First Day
1st Cell Feed. 75 1
3 Hrs.
1st Cell Raffinate
2nd Cell Feed
5 Hrs. (end of run)
4th Cell Raf finale
Second Day
1st Cell Feed. 75 1
1 Hr.
1st Cell Raffinate
2nd Cell Feed
4th Cell Raffinate
3 Hrs.
1st Cell Raffinate
2nd Cell Feed
4th Cell Raf finale
5 Hrs. (end of run)
1st Cell R-f finale
2nd Cell Fscd
LOU
IRON FLOWSHEET. FOUR DAY CONTINUOUS TEST
Concenlralion, gpl
Fe
1.164
0.137
«ne7
1.532
0.265
OQ7
0.237
0.227
0.008
0.249
O79~
Cu
0.069
0.126
O.U08
0.040
0.308
0.28)
0.210
0.105
0.423
0.308
0.210
0.414
0.380
Zn
1.815
1.347
O35
0.014
2.208
' 1.479
O87
0.032
1.433
1.287
0.054
1.304
mz
Cr
5.
5.
5.
5.
4.
4.
4.
5.
3.
4.
4.
3.
4.
749
589
540
808
395
066
957
572
886
213
591
688
031
Nt
5.340
5.318
4.806
3.789
11.511
10.717
8.645
6.298
13.531
11.011
9.739
12 641
12.448
Al
0.041
0.038
0.024
0.009
0.040
0.041
0.025
0.044
0.038
• •••
0.049
0.057
Ca
U.602
0.608
0.636
0.408
0.246
0.229
0.252
1.015
0.158
0.209
0.026
0.151
0.219
pll
1.S8
1.15
1.39
1.19
1.05
1.64
1.50
1.96
1.28
_ T.°C
76
134
36
30
30
SO
30
59
30
-------
TABLE 8.49.
Sample
3331
3351
3353
3354
3J55
335?
3350
3359
3363
3368
3364
3365
3366
3367
Conditions
CONTINUED
Concentration, gpl
4th Cell Raffinate
4th Cell Composite
Raffinate
Third Pay
1st Cell Feed, 75 1
1 Hr.
1st Cell
2nd Cell
4th Cell
3 Hrs.
1st Cell
2nd Cell
4th Cell
1st Cell
2nd Cell
2nd Cell
3rd Cell
4th Cell
4th Cell
Fourth Day
Raffinate
Feed
Raffinate
Raffinate
Feed
Raffinate
Raffinate
Feed
Raffinate
Raffinate
Raffinate
Caapostle
0.040
0.065
1.891
0.377
O5T
0.020
0.368
OI7
O.C08
0.430
O5T
0.056
OZ5
07B24"
BToTff
Cu
0.284
0.182
0.563
0.567
0.456
0.341
0.573
0.521
0.440
0.622
0.534
0.594
0.608
0.551
0.391
Zn
0.028
0.040
2.220
1.499
17127
0.040
1.441
17*55
0.028
1.787
T7555
0.234
070BZ
0"7B5TF
Cr
4.212
4.927
4.931
4.849
4.278
4.110
4.967
4.737
4.550
5.473
4.939
5.392
5.379
5.132
4.599
HI
11.570
8.443
11.650
12.030
12.370
11. b it
12.140
12.196
12.406
13.214
11.972
13.361
13.415
12.996
12.011
Al
0.018
0.011
0.071
0.054
0.058
0.010
0.061
0.056
0.008
0.069
0.061
0.025
0.023
0.019
O.C92
Ca
0.045
0.057
0.372
0.288
0.242
0.024
0.302
0.288
0.052
0.302
0.305
0.199
0.106
0.062
U.04C
1.27
1.11
2.11
1.37
1.08
2.35
1.38
1.05
1.94
1.53
1.22
1.26
T.°C
23
25
54
30
25
50
35
25
41
30
26
3414
1st Cell Feed. 140 1
2.362 0.370 2.436 4.258 8.527 0.057 0.441 1.39 23
-------
TABLE b.49. CONTINUED
Sauple
4415
3421
1417
3418
3419
J422
3427
3423
3424
3425
3428
34)3
3429
3430
3431
3437
3442
3438
3439
3440
Conditions
1 Hr.
1st Cell
2nd Cell
2nd Cell
3rd Cell
4th Tell
3 Mrs.
1st Cell
2nd Cull
2nd Cell
3rd Cell
4th Cell
5 Mrs.
1st Cell
2nd Cell
2nd Cell
3rd Cell
4th Cell
7 Hrs.
1st Cell
2nd Cell
2nd Cell
3rd Cull
4th Cell
Concentration, gpl
Rafflsm
Feed
ftaf finale
Rafflnate
Raffinate
Rafflnate
Feed
Raiflnate
Rattmale
Raffl rule .
Rafflnate
Feed
Rafflnate
Rafftnate
Rafflnate
Rarfinato
Feed
Rat finale
Raf finale
Rafftnate
Fo
0.47S
O55
0.107
O35
025
0.467
insa
0.070
O»
OT027
6.429
O25
0.084
O5S
O6T
0.4«6
OZ5
0.132
OSS
OT7
Cu
0.363
0.513
0.538
O.S99
0.531
0.367
0.423
0.418
0.434
0.423
0.373
0.402
0.391
0.409
0.395
0.379
0.383
0.365
0.400
0.35B
Zn
1.604
rws5
0.304
ITTT?
or?
1.543
T3Z5
0.154
OW
OW
1.684
raff
0.229
0.086
ircss
1.498
T7733
0.194
O79~
O50"
Cr
4.181
4.938
5.072
5.107
5.051
.237
.698
.404
.430
.406
.426
.617
.477
.441
.428
4.439
4.264
4.243
4.356
4.360
Nl
B.593
11.348
12.102
12.544
12.536
8.709
9.880
9.619
9.824
9.915
9.229
9.480
9.420
9.405
9.428
8.795
8.822
8.574
8.713
8.659
Al
0.0522
0.0575
0.0217
0.0160
0.0153
0.0507
0.0495
0.0123
0.0063
0.0099
0.0490
0.0546
0.0144
0.0066
0.0073
0.0526
0.0538
0.0152
0.0095
0.0056
Ca
0.3119
0.3044
0.2711
0.1 503
0.0859
0.3035
0.3051
0.2402
0.1606
0.1197
0.32/4
0.3252
0.2865
0.2320
0.1822
0.3224
0.3306
0.2993
0.2811
0.1911
PH
.20
.76
.57
.49
.12
.20
.88
.67
.56
.27
.22
.73
.59
.48
.27
.21
.95
.64
.54
.31
T.°C
25
40
35
29
26
27
48
43
38
34
29
42
41
36
33
29
48
45
41
36
-------
TABLE 8.49. CONTINUED
Sample Conditions
Fe
ConcenlraHon. gpl
Cu
ro
m
Ot>
? Mrs, (end of run)
3444 1st Cell Raffinate
3453 2nd Cell reed
344b 2nd Cell Raffinale
3446 3rd Cell Raffinale
3447 4th Cell Rafftnate
3454 End of run Composite
Raffinate
H61T51*leach:All leaches performed under standard conditions. 12,000 9 jludge, 1650 cc concentrated
\ Atl *. 9«en •!!' n oo£ *. r..cA
-------
TABLE ft.SO. CELL EFFICIENCY FOR IRON EXTRACTION BY DEKPA: LARGE SCALE CONTINUOUS TESP.10RK
Sample No. Conditions Cell Efficiency, gpl In Rafflnate. I Extracted
—* Total
Cell 1 Cell 2 Cell 3 Cell 4 Efficiency
gpl _* gpl _t gpl % • gpl _*
First Day. 75 1
3281-B First Cell Feed. 1.164 gpl Fe
3Z83-B Second Cell Feed. 0.187 gpl Fe
3284 5 Mrs. (E.O.R.) 0.014 98.8
^ Second Day. 71 1
** 3311 First Cell Feed. 1.532 gpl Fe
3327 Second Cell Feed. 0.279 gpl Fe
3326.28 5 Hrs. (E.O.R) 0.249 83.7 0.040 97.4
3331 Final Composite Rafflnate:
0.065 gpl Fe
Third Day. 75 1
3351 First Cell Feed. 1.891 gpl Fe
3368 Second Cell Feed. 0.351 gpl Fe
3363.64 5 Hrs. (E.O.R.) 0.430 77.3 0.056 84.0 0.026 53.6 0.024 7.7 98.7
05,66
3370 Final Composite Raffinale:
0.095 gpl Fe
Fourth Pay. 75 1
3414 First Cell Feed. 2.362 gpl Fe
-------
TABLE 8.50. CONT!NUEO
Sample No. Conditions Cell Efficiency, gpl In Rafflnate. % Extracted
Cell 1 Cell 2 Cell 3 Cell 4 Efficiency
gpl _l gpl _l gpl Jt gpl _»
3416,17 1 Hr. 0.475 79.9 0.107 76.7 0.045 57.9 0.025 44.4 98.9
18.19
3421 Second Cell Feed. 0.45 gpl Fe
3422,23 3 Mrs. 0.467 80.2 0.070 84.7 0.034 51.4 0.024 29.4 99.0
24,25
g 3427 Second Cell Feed. 0.458 gpl Fe
3428.29 5 Mrs. 0.429 81.8 0.084 80.6 0.068 19.0 0.064 6.2 97.3
30,31
3433 Second Cell Feed, 0.432 gpl Fe
3437.38 7 Hrs. 0.446 81.1 0.132 74.9 0.053 59.8 0.047 11.3 98.0
39.40 ~
3442 Second Cell Feed, 0.526 gpl Fe
3444,45 9 Hrs. (E.O.R.) 0.403 82.9 0.110 77.3 0.051 53.6 0.052 97.7
46,47
345J Second Cell Feed. 0.484 gpl Fe
3454 Final RafMnatc: 0.053 gpl Fe
NOTES:'Detailed data presented In Table 8.49.
'Cell efficiency decreases from cell to ceM because pH of the leach solution drops off.
-------
TABLE 8.51. CELL EFFICIENCY FOR ZINC EXTRACTION BY OEIIPA: LARGE SCALE CONTINUOUS TESTHORK '
4
Staple No. Conditions Stage Efficiency, .jpl In Raffinate. * Extracted j
TSUI
Cell 1 : Cell 2 Cell 3 Cell 4 Efficiency '
gpl _» gpl _l gpl _» gpl _%
First Day. 75 1
32B1-B First Cell Feed. 1.815 gpl Zn
3283-B Second Cell Feed. 2.398 gpl Zn
3284 Five Hours (E.O.R.) - 0.014 99.2
Second Day. 75 1
3311 First Cell Feed. 2.208 gpl Zn
3327 Second Cell Feed. 1.342 gpl Zn
3326. 28 Five Hours (E.O.R.) 1.304 40.9 - 0.028 98.7
3331 Final Cozpostte Raffinate:
0.040 gpl Zn
Third Pay. 75 1
3351 First Cell Feed. 2.220 gpl Zn
3368 Second Cell Feed. 1.558 gpl Zn
3363. Five Hrs. (E.O.R.) 1.787 19.5 0.234 87.0 0.082 64.5 0.060 2C.8 97.3
64. 65.
66
3370 Final Composite Raffinate:
0.043 gpl In
Fourth Day. 140 1
3414 First Cell Feed, 2.436 gpl Zn
3416. 1 Hr. 1.604 34.2 0.304 81.0 0.112 63.2 0.073 34.8* 9/.0
17. 18.
19
-------
TABLE 8.51. CONTINUED
Sap-pie No. Conditions Stage Efflclcnc/. gpl In Raffinale, t Extracted
Total
Cell I Cell 2 Cell 3 Cell 4 Efficiency
3421 Second Cell Feed. 1.605 gpl Zn
3422. 3 Mrs. 1.543 36.7 0.154 90.7 0.066 57.1 0.044 33.3 98.2
23. 24.
25
3427 Second Cell Feed, 1.666 gpl Zn
3428, 5 Mrs. 1.6)54 30.9 0.229 86.0 8.080 65.1 0.060 25.0 97.5
29. 30.
31
N 3433 Second Cell Feed. 1.638 gpl Zn
™ 3437, 7 Hrs. 1.498 38.5 0.194 88.7 0.079 59.3 0.050 36.7 97.9
38. 39.
40
3442 Second Cell Feed. 1.723 gpl Zn
3444, 9 Hrs. (E.O.R.) 1.694 30.5 0.186 88.7 0.068 63.4 0.053 23.9 97.6
45, 46.
47
34S3 Second Coll Feed, 1.644 gpl Zn
34!>4 Final Composite Raffinate:
0.061 gpl Zn
NOUS: 'Detailed data presented in Ubie 8.49.
'Cell efficiency decreases from cell to cell because pH of the cell feed drops off. '
-------
TABLE B.5Z. CELL EFFICIENCY FoH Z1KC PLUS IRON EXTRACTION BY OCHPA: LARGE SCALE CONTINUOUS TESTUORK
at
u>
Sample No. Conditions
First Day. 75 1
5 Hrs. (E.O.R.)
Second Day. 75 >
5 Hrs. (E.O.R.)
Third Day. 75 1
5 Hrs. (E.O.R.)
Final Composite Rafftnate:
0.138 gpl (Fe » Znl
Fourth Day. 140 1
9 Hrs. (E.O.R.)
Final Composite Raffinate:
0.111 gpl (Fe * in)
Cell Efficiency, gpl In Raffinate. t Extracted
Total
Cell 1 Cell Z Cell 3 Cell 4 Efficiency
gpl _• gpl _l gpl t gpl t
0.014
2.SS3 30.5
0.068
99.2
97.0
2.217 46.1 0.290 84.8 0.108 62.8 0.084 22.2 98.0
2.097 56.3 0.296 86.1 0.119 60.0 0.105 11.8 95.1
NOTES: Detailed data presented in Table 8.49.
'Cell efficiency decreases Iron cell to cell because pH of the leech solution drops off.
-------
TABLE 8.53. CONTINUOUS OEHPA TESTKORK DATA SUMMARY: SELL SYSTEM
ro
Sample !!o.
3745
3786
3795
3796
3797
3798
3787
3799
3800
3801
3802
3803
3804
3805
3006
3824
3825
3826
Conditions, End of Run Results Concentration In Final Composite RafHnate. gpl
First Day. 19 1. 14 hrs.
First Stage Feed
Second Stage Feed
Loading
1st Cell Rafflnate
2nd Cell Rafflnate
3rd Cell Rafflnate
4th Cell Rafftnute
Composite Rafflnate
Stripping
5th Cell. H,SO.
6th Cell. '" *
7th Cell, "
8th Cell, HC1
9th Cell, "
10U. Cell. *
Second Day. 19 1, 9 hrs.
1st Stage Feed
2nd Stage Feed
Loading
1st Cell Rafflnate
2nd Cell Rafflnate
3rd Cell Rafflnate
Fe
2.023
TT23S
0.925
C.423
OD5
0.342
BTTSff
0.077
0.075
0.063
3.354
3.496
4.046
2.276
1.364
•
0.515
(T234~
071TC
Cu
0.015
0.017
0.011
0.010
0.018
0.005
0.008
0.006
0.017
0.002
0.005
0.015
0.023
0.021
0.019
Zn
1.628
THM
1.207
OTT
0.043
0.017
OBU
39.35
39.06
38.50
0.403
0.396
0.451
0.354
0.987
0.329
O?ff
onis
Cr
4.620
4.727
4.702
4.632
4.808
4.724
4.679
0.235
0.174
0.167
0.065
0.073
0.083
5.471
5.088
5.434
5.187
f.072
Ml
7.773
7.986
7.853
7.705
8.235
8.124
7.990
0.283
0.2?3
0.222
0.101
0.053
0.070
8.298
8.585
8.234
8.690
8.629
Al
C.033
0.030
0.024
0.002
0.003
0.045
0.037
0.033
0.014
0.025
0.006
0.013
0.044
0.010
0.031
0.026
Ca
0.206
0.147
0.172
G.030
0.009
O.C04
0.015
0.514
0.416
0.415
0.005
0.005
0.028
0.049
0.147
0.046
0.153
.0.135
-------
TABLE 8.53. CONTINUED
s
Ul
Sanple No.
3827
3835
3U28
3829
3830
3831
3832
3833
3846
3847
3860
3861
3862
3863
3871
3864
3865
3866
3867
3868
3869
Conditions. End of Run Results Concentration In Final Composite Raffinate. gpl
4th Cell Rafflnate
Composite Rafflnate
Stripping
5th Cell, H,SO.
6th Cell. '" *
7 tli Cell. "
8th Cell. IIC1
9th Cell. "
10th Cell. "
Third Day. 19 1. 6.5 hrs.
1st Stage Feed
2nd Stage FeeJ
Loading
1st Cell Rafflnate
2nd Cell Raffinate
3rd Cell Rafflnate
4th Cell Raffinate
Composite Rafftnatc
Stripping
5th Cel i . H.SO.
6th Cell. '• *
7th Cell. *
8
-------
TABLE 8.53. CONTINUES
ro
Sample No.
3881
3882
3S95
3896
3897
3898
' syoo
3899
3900
3901
3902
3903
3904
JV'26
3927
3933
3934
3935
Conditions. End of Run Result'.
Fourth Day. 19 1. H hrs.
1st Stage Feed
2nd Stage Feod
Loading
1st Cell Rafflnate
2nd Cell Rafflnate
3rd Cell Raffinate
4tn Cell Rafflnate
Composite Rafflnate
Stripping
5th Cell. H2S04
6th Cell. "
7th Cell, "
8th Cell. HC1
9lhCell. "
10th Cell, '
Fifth Day. 19 1. C.5 hrs.
1st Stage Feed
2nd Stage teed
Loading
1st Cell Raffinate
2nd Cell Raff Irate
3rd Cell Rafflnate
Fe
2.035
F.3TT
0.138
5777?
OTSUo
075TT
OTJI?
0 057
0.069
0.077
3.039
3.352
5.191
2.218
DT509
1.119
OW
065"
Concentration in Final Composite
Cu
0.099
0.016
0
0
0
0
0
0
0
0
0
0
0
.004
.019
.028
.023
.017
.028
.022
.014
...
...
—
...
...
.006
.016
.013
Zn
Cr
2.128 5.304
TrSDT 5.013
1.4?7
T727U
OB7
ITU38"
OT5
38.83
37.27
37.19
0.203
0.182
0.35*
1.999
TUB"
1.921
OTO
OW
5.625
5.026
5.005
4.829
4.839
0.948
0.869
0.904
0.001
^
0.004
5.012
5.613
5.223
5.600
5.729
7
7
7
7
7
0
0
0
0
0
0
6
7
6
8
7
Raffinate. gpl
N<
7.55D 0.
7.269 0.
.64?
.364
.259
.253
.292:
.649
.580
.6U5
.DOS
.004
.013
.615
.867
.962
.005
.956
0.009
0.025
0.012
0.008
—
0.1610
0.1546
0.1446
0.0163
0.0241
0.0267
0.054
0.005
0.050
0.008
0.008
A1
007 0
048 0
O.Olb
0.015
0.008
0.004
0.004
0.8871
0.7150
0.8046
0.0100
0.0133
0.0157
0.003
0.015
0.004
0.009
0.005
Ca
.015
.042
...
...
...
...
—
...
...
...
—
0.055
0.070
_ ••
...
...
-------
TABLE 8.53. CONTINUED
ro
O»
Sample No.
3936
3944
3937
3938
3939
3940
3941
3942
3953
3954
3972
3973
357T
3975
3969
3976
3977
3978
3979
3980
3981
Conditions. End of Run Results
4th Cell Rafflnate
Composite Hat finale
Stripping
5th Cell. H?S04
6th Cell. '
7th Cell. "
8th Cell. HC1
9th Cell. '
10th Cell. *
Sixth Day. 19 1. 6.5 hrs.
1st Stage Feed
2nd Stage Feed
Loading
1st Cell Raffinate
2nd Cell Raffinate
3rd" CTJT Raffinate
4th Cell Raffinate
Cooposlte Raffinate
Stripping
Sth Cell. H2S04
6th Cell. "
7th Cell. •
8th Cell. IIC1
9th Cel 1 . '
10th Cell. "
Fe
0.056
"OZI
0.050
0.067
0.069
2.293
2.506
3.630
2.127
OC5"
0.368
on
...
0.238
OT21B
0.077
0.083
0.092
2.765
3.089
4.350
Concentration In Final Conr.osite Raffinate. opl
0
0
0
0
0
0
0
6
0
0
0
0
0
0
0
0
Cu
.006
.015
.007
...
_-.
...
.020
.020
.027
.024
.028
.026
022
.018
.020
.020
.028
.029
.022
Zn
0.075
ITolT
33.85
32.11
32.06
0.159
0.184
0.274
2.162
I7K7
1.527
IT3B3"
0.108
0.046
OSff
_._
..•
...
0.102
0.117
0.167
Ir N1
5.
5.
0.
0.
0.
.
-
-
5.
4.
4.
•
3.
5.
0.
0.
0.
0.
0.
0.
740
127
766
769
749
._
—
--
712
604
965
319
.•
466
335
769
666
646
018
008
Oil
7.996
7.398
0.532
0.530
0.514
0.010
0.004
0.004
7.124
6.100
6.354
5TBTJ
6.092
4.774
6.763
0.548
0.487
0.487
0.005
0.010
0.014
0.003
0.006
0.153
0.151
0.142
0.013
0.019
0.020
0.010
0.060
0.009
0.038
0.021
0.011
0.012
0.206
0.168
0.175
0.011
0.017
0.018
XI
0.003
0.002
0.6141
0.498
0.508
0.010
0.010
0.010
0.003
0.004
0.005
OD7
0.0016
0.0084
0.0072
0.437
0.389
0.418
0.002
...
0.001
Ca
...
...
...
...
...
0.628
—
0.530
ITJ7T
0.521
0.381
0.517
0.451
0.357
0.480
0.304
0.459
0.546
-------
TABLE 8.53. COMINUED
ro
tn
ao
Sample No,
3992
3993
4UU8
4U09
4010
4011
4022
4012
4013
1014
1015
4016
4017
4U40
4041
4057
4058
4059
Conditions. End of Run Results
Seventh Day. 19 1. 6.5 hrs.
1st Stage Feed
2nd Stage Feed
Loading
1st Cell Raffinate
2nd Cell Raffinate
3rd Cell Raffinate
4th Cell Raffinate
Composite Raffinate
Stripping
5th Cell. HpSOa
6th Cell. "
7th Cell. '
8th Cell. HC1
9th Cell, '
10th Cell. '
Eighth Day. 18 1, 8 hrs.
1st Stage Feed
2nd Stage Feed
Loading
1st Cell Raffinate
2nd Cell Raffinate
3rd Cell Raffinate
Fe
2.299
67557
0.014
OToTT
0.051
0.101
0.088
0.065
1.705
1.U36
2.911
2.040
0.870
0.582
U.1JJ
057
Concentration in Final Composite Raffinate. gpl
LCU_
0.013
0.031
0.313
0.041
0.069
0.046
0 040
0.002
0.016
0.001
0.012
0.012
0.019
0.019
0.015
0.017
0.019
0.025
Zn
2.084
TTTT?
2.122
O55
07135
070T9"
07055
23.62
21.51
21.46
0.049
0.030
0.065
1.987
T742S
1.067
OTZ3T
07C5T
_ Cr
5.
5.
5.
5.
5.
5.
5.
0.
0.
0.
0.
0.
0.
5.
4.
5.
5.
4.
510
914
799
799
784
399
719
496
472
468
008
015
001
338
283
528
300
702
6
7
7
7
7
6
7
0
0
0
0
0
0
6
4
6
6
5
HI _
.520
.315
.193
.298
.466
.838
.149
.359
.345
.337
.020
.007
.016
.030
.988
.465
.313
.651
0.013
0.014
0.016
0.008
0.002
0.003
0.003
0.145
0.139
0.143
0.010
0.011
0.012
0.017
0.018
...
A1
0.008
0.028
0.003
0.003
0.002 •
0.001
0.008
0.326
0.316
0.299
0.010
0.010
0.038
0.019
0.010
...
Ca
0.510
0.435
0.465
0.417
0.395
0.387
0.297
0.127
0.127
0.051
0.126
0.530
0.339
0.357
0.307
0.266
-------
TABLE 8.S3. CONTINUED
Saaple Mo. Conditions. End of Run Results
en
• Zn Cr Nl Al ta
0.039
OSff
0.022
0.01S
0.023
on
4.969
5.478
5.798
6.707
0.251
0.463
0.136
0.139
0.134
2.383
2.577
4.616
0.020
0.011
O.OJ6
...
0.015
0.013
...
...
...
0.078
0.084
0.188
0.579
0.572
0.544
O.OJ>
0.024
0.037
0.372
0.372
0.366
0.018
0.018
0.032
0.177
0.195
0.153
0.213
0.210
C.360
HBTlS~:'Test conditions presented* In Table 8.54.
'First day test, Iron not completely oxidized; therefore. Iron levels In final rafflnate was high.
'Second day test, solution not doped with zinc.
'Fourth day test, iron not completely oxidized.
"Sixth day test, iron not completely oxidized.
-------
TABLE 8.54. COKDITIONS FOR DEIIPA CONTINUOUS TESTVORK TABLES
Test Series
First Day
Second Day
Third Day
Fourth Day
Fifth Day
Sixth Day
Seventh Day
Eighth Day
Aqueous Volume
Da/
19
19
19
19
20
19
• IB
IB
Treated, liters
Total
19
3B
5»
76
96
115
133
151
Tine of Exposure. Mrs. Initial pll
Day
14
9
6.5
B
7
7
7.S
B
Total 1st Cell
14 .22
23
29.5
37.5
44.5
51.5
59
67
.20
.20
.06
.10
.10
.10
.22
2nd Cell
1.83
1.92
(1.48 after 1 hr.)
1.B8
1.99
1.86
1.91
..91
1.81
NOTES: Flow pattern presented In Figure B.20.
'40 v/o OEHPA. 60 v/u KERHAC 510.
•0/A • 1 for both load and strip cells.
'Volume of organic in systca • 7,6 Liters.
'Flow rate of organic and aqueous 40-SOcc/aln.
'Temperature: 20-25 C.
-------
TABLE 8.55. CELL EFFICIENCY FOR IRON EXTRACTION BY DEHPA: LONG TERM TEST
Sample No. Conditions Cell Efficiency, gpl In Rafflnate. \ Extracted
Total
Cell I Cell 2 Cell 3 Cell 4 Efficiency
gpl _1 gpl I gpl I ypl I
Second Day Exposure (19 1)
3805.6 Starting solution. 2.276 q«>!.
Fe. 1st cell Feed. vti * 1.20;
2nd Cell Fccu, pll = 1.92 Ini-
tltally for 1 hr. then Ac- *
ouaseo to 1.48. 1.364 gpl Fe
3807.8.5. IB Mrs. 0.400 82.4 0.088 93.S 0.055 37.5 0.055 0 96.0
10 ~
3824.25. 23 llrs. 0.515 77.4 0.234 82.8 0.146 37.6 0.109 25.3 92.0
26.27
3835 End of Run Rafftnate (E.O.R.):
0.070 gpl Fe
Third Pay Exposure (19 I)
3846.47 Starting solution, 2./42 gpl Fe;
1st Cell Fcej, pli - 1.20; 2nd
Cell Feed. 1.88. 0.485 gpl
3848.49. 26 Mrs. 1.208 55.9 0.050 B9.6 0.050 0 0.050 0 89.7
50.51
3860.61. 29 Hrs. 0.854 68.9 0.055 68.9 0.018 67.3 0.027 0 . 94.4
62.63
3871 E.O.R. Raffinate: 0.30 gpl Fe
-------
TABLE 8.S5. CONTINUED
Sanple Mo. Conditions Cell Efficiency, gpl In Rafflnate. % Extracted
Cell I Cell 2 Cell 3 Cell < Efficiency
gpl t gpl I gpl _t gpl _\
Fifth Pay Exposure (20 ])
3926,27 Starting Solution, 2.218 gpl Fe;
1st Cell Feed. pH • 1.10; 2nd
Cell Feed. pH • 1.86. 0.509
gpl Fe
3916,17, 41 Mrs. 1.149 48.1 0.094 81.5 0.060 36.2 0.035 41.7 93.1
18,19
3933.34, 44.5 Hrs. 1.119 49.5 0.114 77.6 0.065 43.0 O.OS6 13.8 89.0
35>36
3944 E.O.R. RafMnate: 0.022 gpl Fe
Seventh Day Exposure (18 1)
3992,93 Starting Solution. 2.299 gpl Fe;
1st Cell Feed, pH • 1.10; 2nd
Cell Feed. pH • 1.91. 0.567
gpl Fe
4008.9. 57.0 Hrs. 0.014 99.4 0.014 0 JO.L. 100 SD.L. 100 100
10,11 — ~ — —
4022 E.O.R. RafHnate: 0.051
Eighth Pay Exposure (IB 1)
4040,41 Starting Solution. 2.040 gpl Fe;
1st Cell Feed. pH • 1.22; 2nd
Cell Feed, pll - 1.81. 0.870
gpl Fe
-------
1ABLE 8.SS. CONTINUED
CM
«J
OJ
Sample Ho. Conditions Coll Efficiency, gpl In Rafflnate. t Extracted
Cell I Cell 2 Cell 3 Cell 4 Efficiency
gpl j* gpl _l gpl t gpl _X
4042.43
44.45
4057.58
S9.60
4054
NOTES:
62 Mrs.
67 Mrs. 0.582
E.O.R. Rafflnate: 0.028 gpl Fe
'Detailed data presented1 in Table 8.54.
0.132
71.5 0.133
64.6 0.046
77.1 0.057
65.2 0.040 13.0
SO. 8 0.03 46.2
95.4
95.5
•pll of feed to cell, adjusted to 1-1.2; Raffinale fed to a separate reservor; pll of feed to Cell 2
readjusted to pll of l.U. solution fed from reservoir to Cell 2. Usually feed for Cell 2 was
rafflnate from previous exposure. t+
•Run No. 1, 4. 6 omitted because iron was not completely oxidized and OEHPA does not extract Fe .
-------
TABLE 8.56. CELL EFFICIENCY FOR UK EXTRACTION BY BE HP A: LONG TERN TEST
Sample No. Conditions
3745, 3789
3775.6.6,8
3795,96.
97.98
3787
3846.47
3848,49.
50,51
3860.61,
62,63
First Day Exposure (15 1) •
Starting Solution. 1.828 gpl Zn
1st Cell Feed, pll « 1.22; 2nd
Cell Feed. pH - 1.81, 1.070
gpl Zn
9 Mrs.
14 Hrs.
End of Run (E.O.R.) Composite
Rafflnate: 0.080 gpl Zn
Third Day Exposure (19 1)
Starting solution. 2.207 gpl Zn
1st Cell Feed, pH - 1.20; 2nd
Cell Feed, pH « 1.68, 0.299
gpl Zn
26 Hrs.
29.5
Cell Efficiency, gpl In Rafflnate, I Extracted
Cell 1 Cell 2 Cell 3 Cell 4
Total
Efficiency
gpl
gpl
gpl
_S2l_
0.912 50.i 0.140 86.9 0.041 70.7 0.032 22.0
1.207 34.0 0.117 89.1 0.043 64.1 0.017 60.5
97.0
98.4
1.641 25.7 0.036 88.0 0.044 0.059
80.3
0.975 55.8 0.043 85.6 0.039 9.3 0.035 10.2 88.3
3871
E.O.R. Rafflnate: 0.106 gpl Zn
-------
TABLE 8.56. CONTINUED
ro
^i
ui
Sample Ho. Conditions
3681.82
3883,4.5.6
3895.6.7.8
3908
3926.27
3916.17.
18,19
3933,34.
35.36
Fourth Day Exposure (19 1)
Starting solution, 2.128 gpl Zn
1st Cell Feed. pH •> 1.06; 2nd
Cell Feed. pH - 1.99; 1.501
gpl Zn
34.0 Hrs.
37.5 Hrs.
E.O.R. Raffinate: 0.046 gpl
Fifth Day Exposure (20 I)
Starting Solution. 1.999 gpl Zn
1st Cell Feed, pll - 1.10; 2nd
Cell Feed, pH - 1.86; 1.636
gpl Zn
41 Mrs.
44.5 Hrs.
Cell Efficiency, gpl in Raffinate. t Extracted
Cell 1 Cell 2 Cell 3 Cell 4
Total
Efficiency
gpl Jl _ gpl _i _ gpl
gpl
1.688 20.6 0.221 85.4 0.061 72.4 0.040 59.5
1.427 32.9 0.270 82.0 0.082 69.3 0.038 53.7
97.3
97.5
1.858 14.1 0.339 79.3 0.102 69.9 0.052 49.0 96.8
1.921 3.9 0.440 73.1 0.129 70.7 0.075 41.9 95.4
3944
E.O.R. Raffinate: 0.032 gpl Zn
-------
TABLE 8.56. CONTINUED
Sample No. Conditions Cell Efficiency, gpl in Rafflnate. t Extracted
Total
Cell 1 Cell 2 Cell 3 Cell 4 Efficiency
Sixth Day Exposure (19 1) gpl I gpl % gpl » 9Pl *
3953,54 Starting solution, 2.162 gpl Zn
1st Cell Feed. pH - 1.10; 2nd
Cell Feed, pH - 1.91; 1.617
gpl Zn
3956.57. 50.5 Mrs. 1.868 13.6 0.141 91.2 0.141 0 0.054 61.7 96.7
5B.59 ~
SJ 3972.73. 51.5 Mrs. 1.527 29.4 0.383 76.3 0.108 71.8 0.046 57.4 97.2
74.75
3975 E.O.R. Rafflnate: 0.050 qpl Zn
Seventh Day Exposure (IB T)
3992.93 Starting Solution, 2.084 gpl Zn
1st Cell Feed. pH • 1.10; 2nd
Cell Feed, pH • 1.91; 1.779
gpl Zn
4008.9.10 57 Hours 2.122 0 0.455 74.4 0.13-J 70.3 0.039 71.1 97.8
11 ~ ~~
4022 E.O.R. Rafflnate: 0.066 gpl Zn
Eighth Day Exposure (IB 11
40411,41 Starting solution, 1.9B7 gpl Zn
1st Cell Feed, pll - 1.22; 2nd
Cell Feed, pll - 1.81; 1.420
ypl Zn
-------
TABLE 8.56. CONTINUED
Sanple No. Conditions Cell Efficiency, gpl In Raffinate. 1 Extracted
Total
Cell I Cell 2 Cell 3 Cell 4 Efficiency
gpl _t gpl _X gpl _1 gpl _»
4042.43. 62 Hrs. 0.209 65.3 0.03B 81.9 0.025 .34.2 98.2
44.45
4057.58 67 Hrs. 1.067 46.3 0.231 83.7 0.061 73.6 0.023 62.3 98.4
59.60
4054 E.O.R. Reffinale: 0.043 gpl Zn
NOTES:'Detailed data presented In Table 8.54.
•pll of feed to Cell 1 adjusted to 1-1.2; raffinate fed to a separate reservoir; pll of feed to Cell
2 readjusted to 1.8.
•Run No. 2 omitted because zinc content very low.
-------
TABLE 8.57.CELL EFFICIENCY FOR IRON AMD ZINC EXTRACTION BY UEHPA: LONG TERM TEST
Sanplc No. Conditions
ro
>j
GO
3848,49
50.51
3860.61
62.63
3871
3916,17.
18.19
3933.34
35.36
3944
400S.9.
10.11
4022
(See Tables 8.55, 8.(6)
Third Day Exposure
26 Mrs.
29.S Mrs.
End of Run it.O.K.) Rafflnate:
0.137 gpl Fe * In
Fifth Day Exposure
41 Hrs.
44.5 Hrs.
E.O.R Rafflnate: 0.053 op]
Fe_*_Zn
Seventh Day Exposure
57 Mrs.
E.O.R. Raffinate: 0 117 gpl
Fe * Zn ~~
Cell Efficiency, gpl to Rafflnate. % Extracted
Cell 1 Cell 2 Cell 3 Cell 4 Efficiency
gpl « gpl _* qpl _% gpl t
2.849 42.4 0.086 89.0 0.090 •--- 0.109 8C.1
1.829 63.0 O.C98 07.5 0.057 41.8 0.062 92.1
2.197 47.9 0.433 79.8 0.162 62.8 0.087 46.3 95.9
3.040 27.9 O.b54 74.2 0.194 35.0 0.131 32.5 93.9
2.136 51.3 0.469 80.0 0.135 71.2 0.039 71.1 98.3
-------
TABLE 8.57. CONTINUED
IM
Sample No. Conditions
4042.43
44.45
4057.5U
59.60
4054
Eighth Day Exposure
62 Mrs.
67 Mrs.
E.O.R. Raffinate: 0.071 gpl
Fe + Zn
Cell Efficiency, gpl In RftfflnUe. 1 Extracted
Total
Cell I Cell 2 Cell 3 Cell 4 Efficiency
gpl _» gpl _% gpl _1 gpl _»
0.341 85.1 0.084 75.4 O.OC5 22.6 97.2
1.649 59.1 0.341 85.1 0.084 75.4 0.065 22.6 97.2
NOTES:
Tables 8.54. 8.55. and 8.56.
-------
carried over to the other extraction cells and into the strip cells thereby
contanimating the strip solution; and excessive loss of organic occurred.
An analyses of the crud material showed high iron and phosphorus contents.
Therefore, a series of tests were conducted to investigate the role of iron
content in the aqueous phase. The experimental approach consisted of:
'cleaning out the Reister SX testrack; refilling with new 40 v/o
09EHPA. 60 v/o KERMAC 470B. 200 gpl H.SO.. strip acid (3 cells] and 4
ITHC1 strio acid (1 cell). * *
'exposing the new organic and strip solutions to first pure zinc
sulfate solutions; then to a low iron bearing feed; then to a mixed
metal feed. An organic/aqueous ratio of one was maintained in all
extraction and strip cells. (The flow pattern is presented in
Figure 8.15.)
ZINC SULFATE TEST RESULTS
Conditions: 75 liters feed solution, 2.46 gpl Zn.
Unused 40 v/o DgEHPA, 60 v/o KERMAC 4708; volume = 32.8 liters.
jsed 200 g/1 H.SO. a
.4 g/1, volume » 12.
Unused 4 N riCl in rack, volume =4.1 liters.
Cells 1, 2, 3, 4 feed rate maintained @ 0.25 1/min.
Cell 1 feed @ pH = 2.0
Cell 3 feed G> pH = 2.0. adjusted with 500 g/1 KOH.
Strip acid was replaced hourly to maintain acid concentration at
200 gpl.
Method: Evaluation of crud formation was strictly by observation with most
attention given to cell #1 settler. Crud formation was judged
critical when:
1. interface level in cell $1 became uncontrollable, caused by crud
clogging aqueous jack-leg line.
2. crud overflowed organic weir for transport to strip section with
subsequent contamination of strip acid.
Results: Chemical results for the test series are presented in Table 8.58.
280
Unused 200 g/1 H.SO. acid doped with ZnSO. to Zn concentration
30.4 g/1, volume = 12.0 liters. *
-------
TABLE 8.58. ZINC SULFATE SOLUTION CRUD FORMATION TEST: Zn
RECOVERY
Condition Zinc Concentration In Aqueous Phase (gpl)
Cell 2 Cell 4
Starting Solution, 75 liters
2.46 gpl Zn, pH = 1.96.
Rafflnate
45 m1n.
2 Hrs
3 Hrs
4 Hrs
0.56 (pH
0.52
0.41 (pH
0.22 (pH
« 1.36)
= 1.37)
- 1.35)
0.02
0.01
Notes: . Test conducted In large scale Reister SX testrack.
. See text for test conditions.
281
-------
Comments: First stage extraction cell #1 showed considerable crud formation
approximately half-way through the run; temperature of feed
solution was at 22.0 C. Feed solution was heated to approximately
40 C, and although crud did not dissipate, it did stop increasing
in amount.
Strip acid replacement was performed entirely in strip cell *3
settler. Evaluation of Zn concentration in each strip cell at end
of run showed significant differences in Zn concentration between
cells, which was assumed to correspond inversely to H-SO.
concentrations. Strip acid flow rates between cells wert
approximately 0.2 liters/min. allowing for approximately two cell
volumes of aqueous oer hour through each cell. Obviously this was
insufficient volume change to distribute new acid throughout the
strip system each hour. Based on this observation, acid
replacements in all subsequent experiments was performed by equal
volume replacement in each strip cell settler. Strip acid
replacement calculation was based on 3.0 g/1 Zn in feed solution
when actual Zn concentration in feed was 2.46 g/1. 'this resulted
in over-dilution of Zn concentration in strip acid through course
of this experiment. End of run data indicated Zn average
concentration to be 24.7 g/1 instead of projected 30.4 g/1.
MIXED METAL FEED SOLUTION TEST RESULTS
Purpose of Experiment:
Provide mixed metal feed solution to ZnSX system and '-.onitor crud
formation. Evaluate effectiveness of equal volume strip acid
replacement in each strip cell settler.
Conditions: 75 liters mixed metal leach solution produced from Morris
Industries sludye; i.e., 4470 g sludge, 740ml H.SO., diluted to
75 liters with 01 water.
0-EHPA from previous test.
H,SOa strip solution left from previous test; 24.75 gpl Zn; volu-ne
=*12?0 liters, three H2$04 strip cells.
HC1 volume Increased for this run by placement of an external
reservoir of 4N KC1, total volume HC1 = 19.1 liters, one HC1 strip
cell.
Cells 1. 2 ,3. 4 feed rate @ 0.25 liters/min.
Cells 1 and 2 feed pH = 2.0.
Cells 3 and 4 feed pH = 2.0, adjusted with 500 g/1 KOH.
Method: Evaluation of crud formation same as for zinc sol fate solution
test.
-------
Results: Chemical results for the test series are presented In Table 8.59.
Comments: Crud formation with this feed was more severe than that
encountered with the previous run. Crud in cell rfl settler was
removed after 4 1/2 hours of system operation (i.e., 90% of feed
volume through cell #1) because crud began overflowing organic
weir.
Evaluation of strip section, cell-by-cell Zr. concentrations showed
equal volume acid replacements in each settler to be effective in
reducing differences In Zn concentrations between cells.
Assay of HC1 strip, showed practically no Fe content at end of
experiment; also HC1 strip showed no yellow color characteristic
of even low concentraticr. of Fed.. Conclusion drawn was that HC1
used to mix strip had previously Seen diluted, resulting in a
strip solution far below 4 N in HCi concentration.
NIXED METAL FEED SOLUTION TEST RESULTS. HIGH IRON
Purpose of Experiment:
Test crud formation 1n ZnSX system with a high Iron bearing,
mixed-metal feed solution.
Conditions: 75 liters mixed-metal leach solution produced from Morris
Industries sludge; i.e.. 4470 g sludge, 740 ml H-SO., diluted to
75 liters with DI water. Ferrous sulfate added Co oring^Ee
content of feed up to 3.0 g/1. H,0, added to oxidize Fe to
Fe 3. fc *
Method:
Results:
Comments:
D-EHPA from previous test.
H,SO. strip replacement scheme as before, volume of strip
ifters.
12.0
HC1 from previous test replaced with 4 N HC1 solution made from
reagent grade HC1, 18.0 liters in system, one HC1 strip cell with
external reservoir.
Cells 1, 2, 3, 4 feed rate @ 0.25 liters/min.
Cells 1 and 2 feed 9 pH » 2.0.
Cells 3 and 4 feed 9 pH • 2.0, adjusted with 500 g/1 KOH.
Evaluation of crud formation same as for previous test.
Chemical results for the test series are presented in Table 6.60.
Crud formation for this run was extreme. 'Cell 91 of first
extraction stage became uncontrollable after 35 liters of feed
283
-------
TABLE 8.59. MIXED METAL SOLUTION CRUD FORMATION TEST: IRON AND ZINC
RECOVERY
Conditions Metal Concentration in Aqueous Phase (gpl)
Cell 2 Cell 4
Starting Solution, 75 liters
0.33 gpl Fe, 2.74 gpl Zn.
1.80 gpl Cr, pH = 1.95.
Raffinate
l.b Mrs
2.5
a.O
5.0
Fe
-------
TABLE 8.60. HIGH IRON MIXED METAL SOLUTION CRUD FORMATION TEST:
IRON AND ZINC RECOVERY
Conditions Metal Concentration in Aqueous Phase (gpl)
Cell 2 Cell 4
Starting Solution, 75 liters
2.71 gpl Fe. 2.68 gpl Zn,
1.63 gpl Cr, pH = 2.03
Raffinate
1.
3.
5 Mrs
0
Fe
0.01
(pH -
-------
because jack-leg plugged with crud. Experiment was halted, cell
#1 settler cleaned of crud, and temperature of the feed increased
from 20 C to 45 C. Operation continued with only a slight
decrease in crud format-Ion rate.
Cell fl'of the first stage extraction appeared to be removing a
good portion of the Fe present in the feed solution. Color change
between cell 91 feed and cell #1 raffinate led to this suspicion.
The feed stream was dark green in color; characteristic of a high
Fe bearing solution, and the raffinate stream from cell #1 was
light blue in color: characteristic of the Norris sludge leach
solution before Fe addition.
8.4.3.2. Crud Problem Solution
The crud formation problem was overcome by switching the kerosene diluent.
KERMAC 510 was substituted for the previously used KERP.AC 470B. A comparative
analysis between the two kerosenes is presented in Table 8.61. The major
difference is the aromatic content, i.e., 4708 has a nominal 11.7* aromatic
content, 510 has a nominal 2.5» aromatic content.
All subsequent testwork was performed using 510 as the diluent. The
long-term continuous testwork was initiated using 470S. Crud formation was
initially a problem in that test set-up (Bell Engineering Testrack) but the
problem disappeared when the diluent was switched from 470B to 510. Phase
separation, metal value selectivity, metal value recovery (Zn plus Fe), and
interface control were excellent throughout the test series.
8.5. SOLID-LIQUID SEPARATION
An Ingersol-Rand 360 IX LASTA pilot scale filter press (shown pictorally
in section 8.14) was purchased for the project. The particular press system
was chosen so that a number of experimental variables could be investigated,
e.g., cake compression, wash options, flow rates, back pressures, temperature,
filter cloth porosity, air drying. Extensive investigation to establish
optimum filtering conditions has not been performed. Filter press features are
presented in Table 8.62.
The filter press has been useci on large scale tests to separate the
jarosite-leach residue solid mixture from the solution. Tests have been
performed on filterability of leach residue and on mixtures of jarosite-leach
residues. Typical results are summarized in Table 8.63.
236
-------
TABLE 8.61. COMPOSITION OF DILUENT USED FOR DISSOLUTION OF DEHPA
Diluent Composition. %
Paraffins Naphthalenes Aromatics
KERMAC 4703 48.6 39.7 11.7
KERMAC 510 2.5
NOTES: 'Composition was not determined during this study. Values reported are
from literature sources (Ref. 28).
TABLE 8.62. PILOT SCALE IR LASTA FILTER PRESS
Material of Construction: All wetted parts 316SS except dlaphrams are
natural rubber
Filtering Area: 0.104 H2 (2.2 ft.2)
Chamber Volume: 1.2 liters (0.3 gal.)
Pressure: 9
Feed Solution 7 kg/cm%(99.S psi)
Compression water 15 kg/cm (213 psi)
for diaphragm testing. to
20 kg/cm' (284 psi)
The filtering rate for a jarosite solid Is much greater than the filtering
rate of the leach residue. This Is, In fact, one of the main reasons for
removing Iron as a jarosite, i.e., the ease in solid/liquid separation.
The jarosite filtering rates achieved in this study are compared to
commercial filtration data supplied by Ingersol-Rand, Table 8.64.
287
-------
TABLE 3.63. FILTERABILITY TESTUORK
Leach Residue; Test 29-2
Operating Sequence Time !min.)
Feed
Core Blow
Top Wash
Precompression
Backwash
Compression
7
0.3
3
1
3
5
Volume Liquid
Recovered (liters)
3.2
1.7
0.9
1.1
0.5
Suspended
Solid (rng/1)
75
)
) Composite 1
Resulting cake moisture (average of three separate pre:.; tests): £6.£%. so
Solid 1ceding rate (average of three separate press tests): 4.5 kg/nr/br.
solids.
Jarosite-Leach Residue Mixture
Feed Test No.
5-1
5-2
5-3
5-4
5-5
Top Wash Test.
5-1
5-2
5-3
5-4
5-5
Time (nrin.)
6
8
13
11
9
4
4
4
4
Volume Liquid
Recovered (liters)
20.1
39.1
20.6
15.9
13.9
18.3
3.9
3.0
2.3
2.4
Suspended
Solid (mg/1)
Resulting Cake Moisture and Loading Rate
Test No. % Solids
5-1
5-2
5-3
5-4
5-5
67.9
68.6
71.2
66.9
66.2
Loading Rate
kg/ir.2/hr.
54.0
41.0
26.3
29.1
35.1
288
-------
TABLE 8.64. JAMSITE FILTRATION RATES
Present Study Japan* Canada*
Feed Solids ' . 40-50 50-55 20-30
Solids Loading Rate 25-55 80-100 40-50
(kg/mZ/hr.)
Cake Solids (%) 66-71 78-80 75-78
*data supplied by Ingersoll-Rand(lt6)
289
-------
Settling rates are exceptionally rapid for jarosite solids. The
experimental solid/iiquid separation in the large scale testwork was. in fact,
accomplished by allowing the solids to settle; pumping a major portion of the
liquid out of the leach vessel; then filter pressing the remaining slurry.
These tests were performed in a 270 liter vessel containing about 200 liters of
solution plus solids. The jarosite settling was essentially complete in less
than 30 minutes.
8.6. COPPER SOLVENT EXTRACTION TESTWCRK
8.6.1. Small Scale Preliminary Testwork
A large number of preliminary small scale shake tests were performed to
characterize several commercially used copper extractant reagents, LIX-64N,
LIX-622.
8.6.1.1. LIX 64N
Preliminary shake tests were conducted to establish potentially important
variables and experimental procedure. The design matrix that was developed
from preliminary tests is presented in Table 8.65. The results show that
copper extraction is primarily a function of LIX-64N content, and leac>-
solution pH and that the presence of deconol decreases copper extraction. The
other variables have minimal influence on copper extraction. The effects
portion of the table indicates that none of the six variables studied greatly
influence the extraction of other elements. Conditions can be chosen from the
design table matrix to achieve effective and selective copoer extraction.
A number of other tests were conducted using the LIX-64 reagent. The
influence of pH on copper extraction (other metal extractions are shown in a
data table on the same figure) is presented in Figure 8.16. Copper is
preferentially extracted from the leach solution at pH values up to 1.75, I.e.,
Cr, Mi, Zn, Fe, and Cd are not extracted. Some iron was extracted at a pH of 2
and above. Note, however, that in these tests there was a large excess of
LIX-64 reagent above what is required to remove the low copper content, i.e..
further experimental work showed that approximately 0.1 gpl Cu is removed from
290
-------
ro
10
TABLE 8.65. DESIGN MATRIX FOR LIX 64N EXTRACTION OF COPPER FORM SLUDGE LEACH SOLUTION (1/8 REPLICA)
s
•
•
219
280
?BI
KJ
?H-I
?Bt
m
786
288-
a»i
Unit
High (.)
ic. (-)
Itil «
0 Outline
Effects
Cu
Fe
Cr
Kt
Zn
Cd
III Ul
«)
15
5
20
10
-
*
•
t
-
»
•
+
10.4
0.8
0.6
0.4
- Z.8
- 0.2
l»pi
Iiroitnt
47QB
--
470B
'50
-
-
i
t
-
-
t
+
1.4
0.9
0.7
-0.4
-7.4
0.5
Otconol
(X)
10
10
20
0
-
-
-
-
t
*
*
»
-26.9
- 1.0
- 0.8
- 0.7
-4.4
-o.z
Hit
(•in.)
3
2
5
1
-
»
»
-
»
-
-
4
S.9
0.1
-0.9
0.7
-5 1
•O.Z
Itip.
!D0
40
15
55
25
-
-
*
*
t
t
-
.
2.9
-Z.5
-l.Z
-Z.4
-3.3
-1.7
I tacb
Solution
pH
1.75
0.25
2.0
1.5
-
t
-
»
»
-
*
_
17,1
5.1
l.Z
1.8
- 3.6
* Z.O
Hli. f«p.
Results - Extraction by Organic (X)
Cu
31.6
»-H
il.fi
5,1
t.i(fl.e
4.2
4.7
3,9
9.J14I
»9.1
Fe
10.5
17.5
2.5
19.0
M.« (M.»)
0
15.5
12.0
6.S • 1.0
*?.6
Cr
1
I
.3
.6
5.8
u.ofu.c
0
5.0
2.5
i.ljl.l
*3.2
Virittion
NOTE: -Leach Sol
•Initial S
Cu 0.45 ^
Zn 1.40 i
Hi 0.86 1
HI •
9.6
IS.fl
2.8
8.5
$.8(1.1)
1.7
8.5
9.6
».ii?.J
±4.7
Zn
42.8
10.0
1.4
8.6
Z. 1(1.1)
12.1
2.1
3.6
8.6:0.8
±2.6
Cd
8.6
7.4
1.2
9.8
1.6(9.6)
0
8.6
8.6
9.8;I.S
i5.6
ution from Sludge barrel 1
olutlon Composition (gpl)
• 0.01. Cr 0.24 10.02.
. 0.03, Fe 2.00 i n.04.
. 0.03. Cd 0.0311 0.002
•Organic/Aqueous • 1: 50cc each
•Test 5 Duplicated
•Baseline Run Three Times
-------
100
80
u
ID
L.
0*
£
o
60
40
20
I
10 i/o Llx 6tN
90 v/o Napoleui UOB
0/A . 1
I »p . Z5°C
Tiie » 3 »in.
eieient Cone. C»tU) Cone. t»t(S) Cone. E«t(*)
Cu
Cr
Ni
Zn
Fe
Cd
0.17 63.8 0.05
0.26
0.82
1.S2
1.97
0.08
0
0
0
0
0
0.27
O.Bt
1.46
2.03
0.08
9.8 0.05 91.*
0.26 0
O.Bt 0
1.16 0
1.5S U.b
0.08 0
0
0
0
0
0
Initial Solution Concentration (gpl)
Cu
Cr
Ni
Zn
Fe
Cd
0.46 ; 0.02
0.27 • 0.02
0.82 1 U.OS
l.tS 1 0.08
1.86 1 0.11
0.08 1 0.005
I
1.50
1.75
Initial pH
2.00
Figure 8.16. Influence of pH on LIX 54N extraction
292
-------
a leach solution by each volume percent (v/o) of LIX-64 reagent; Table 8.66.
This means then, that an excess amount of LIX-64N was present in the Isotherm
tests and that other cations may be picked up (at the hlqher pH levels) once
the copper
-------
ro
£
TABLE 8.66. SUMMARY
Initial pH 10 5
pH * 1.50 0.09
pH • 1.75 0.09 0.10
pH • 3.00 O.M
NOTE: . Maximum copper loading 6.6 gpl.
OF COPPER LOADING IN LI X 64 N
gpl Cu / v/o LIX 64N
0/A
2 1
0.06
0.10 0.08
(0.10)
0.10
(40v/o)-KERMAC 470B
0.5
0.06
(0.06)
^^
0.2
0.02
0.03
0.04
-------
100
80
60
o
TJ
«
L.
**
X
40
20
Conditions: One Contact
Three •inutes
I »p. - 25°C—
Initial Cone. 5 v/o L»
10 v/o IH
IS v/o LIX
15 v/o LU
Eluent Cone.
Cu
Cr
Ni
Zn
_ Cd
Cone. t»t (t) Cone. C.t (*) Conr. tit (*) Cone. Eit (*.)
0.1.6 1 .02 0.17 63.0 0.13 71.1 0.0*6 90.0 O.OSO 89.2
1.78(2.06) 0 1.90(1.97) 0 1.72(2.07) 0 1.67(2.03) 0
0.25 0 0.26( .24) 0 0.2S 0
0.80 0 0.8S 0 0.81 0
1.52 0 1.50 0 1.56 0
1.861 .11
0.27 1 0.02 0.25( .26) 0
0.82 1 0.05 0.80 0
1.4810.08 1.52 0
0.08 1 .005 0.08( .08) 0 0.07(0.08) 0 0.08( .08) 0 0.08(0.03) 0
J_
J_
8 12 16
v/o LIX 64N In KE3MAC 4708
20
24
Figure 8.17. Influence of LIX 64N concentration at pH = 1.76.
2S5 .
-------
2.4
2.0
1.6
g, ,
enl.Z
o
0.8
0.4
T
T
T
Organic: 10 v/o LU 64N. 90 v/» Napoleui 470 B
pH . 1.98
Initial Solution Coip. (gpl)
1.0510.06
1. 7510.10
0.1010.006
(Data presented in Table 8.50)
I _ I _ I
0.05
0.10 0.15
Aqueous (gpl)
0.20
0.25
Figure 8.18. McCabe-Thiele equilibrium extraction isotherm: LIX 64N,
pH = 1.98.
296 .
-------
2.4
2.0
1.6
a.
01
u
•F"
e
a
0.8
0.4
T
T
T
Organic: 10 v/o LIX 64H. 90 v/o Napoleui 4708
pH - 2.20
.05
Initial Solution £oip. (gal):
I
I
0.10 0.15
Aqueous (gpl)
0.20 0.25
Figure 8.19. NcCabe-Thiele equilibrium extraction Isotherm: LIX 64N,
• pH = 2.2.
297
-------
TABLE 8.67. McCABF-THIElE DIAGRAM DATA: pll - 1.98
0/A Ratio
10
5
2
1
O.S
«
0.2
Element Concentration (qpl)
Cu
Organic Aqueous
0.060
0.256 O.or. (0.01B)
0.376 (0.402) 0.035 (0.031)
0.670 (0.670) 0.058 (0.032)
1.004 (0.996) 0.065 (0.074)
1.600 0.175 (0.196)
Aqueous
Fe
Z.26
— (2.16)
2.26
2.19
2.29
2.30
Cr
0.26
0.24 (0.25)
0.26
0.25
0.26
0.26
Nl
0.94
0.97 (0.95)
0.95
0.96
0.94
0.97
In
1.66
1.46 (1.60.»
1.64
1.60 .
1.66
1.70
CJ
0.09
0.09 (0.09)
0.09
0.09
0.09
0.09
ro
vo
00
Starttnq solution (gpi): 0.54:0.04 Cu. 2.46:0.12 Fe. 0.28i0.02 Cr. 1.0Si0.06 Nl. 1.75:0.10 Zn, O.lQi.006 Cd.
-------
TABLE 8.68. HcCABE-THIELE DIAGRAM DATA: pit • 2.20
0/A Ratio
10
S
2
1
0.5
0.2
Element Concentration (gpl)
Cu
Organic
0.143
0.234
0.400 {0.378}
0.596
1.166 (l.Ht)
1.540
Aqueous
0.009
0.005 (0.008)
0.007 (0.007)
0.011
0.049 (0.022)
0.179 to. 176)
Aqueous
Fe
2.18
2.13
2.10 (2.17)
2.23
2.16 (2.19)
2.21
Cr
0.25
O.-M (.25)
0.24 (.25)
0.25
0.25 (.25)
0.25
Ni
0.95
0.86 (0.92)
0.90 (0.94)
0.96
0.91 (0.94)
0.94
Zn
1.62
1.60 (1.58)
1.55 (1.60)
1.65
1.58 (1.60)
1.62
Cd
0.09
0.08 ( .08)
0.08 (0.09)
0.08
0.08 (0.09)
0 09
IM
VO
10
Starting solution (gpl): 0.54»O.C4 Cu. 2.46i0.12 Fc. 0.28:0.02 Cr, 1.05:0.06 Nt, 1.75:0.10 Zn. 0.10i0.006 Cd.
-------
IABLE 8.69. LlX 64N ISOIHERM DATA: 40 v/0 LIX 64N APPLIED 10 LIACH SOLUTION
Sample No.
1061
1064
U
0 1069
1066
1067
1070
1068
1071
1072
1073
Condition
Starting
pH - 1
0/#
0/A
0/A
0/A
0/A
0/A
0/A
pH ° 1
0/A
0/A
0/A
Solution
.75
- 5
• 2
« 2 (repeat)
• 1
- O.S
> 0.5 (repeat)
*»
- 0.2
.5
VIOI^
• \f
• 1
• 0.2
Cu
4.86
0.82
0.90
0.87
1.52
2.34
2.37
3.54
1.22
2.28
3.97
Fe
15.16
14.68
14.32
14.59
14.58
14.97
15.22
15.23
14.64
15.20
15.06
Concentration (gpl)
Zn
10.
10.
9.
9.
9.
10.
10.
10.
9.
10.
10
34
02
76
87
88
16
31
32
92
27
15
Hi
4.12
4.04
3.97
4.06
4.06
4.18
4.24
4.26
4.04
4.23
4.17
•w
1
1
1
1
1
1
1
1
1
1
1
Cr
•^•^^^v
.04
.02
.00
.02
.01
.04
.06
.06 :
.02
.05
.04
Cd
0.58
0.57
0.56
0.57
0.57
0.59
0.60
0.60
0.57-
0.60
0.59
Al
4.71
4.58
4.46
4.53
4.50
4.67
4.72
4.73
4.53
4.66
4.62
-------
TABLE 8.69. CONTINUED
Sample No. Condition
PH • 2.0
1074 0/A • 10
1075 0/A • 1
1076 0/A « 0.2
Concentration (gpl)
Cu fe Zn Ni
0.34 15.38 10.44 4.26
0.94 15.40 10.45 4.38
3.36 16.07 10.81 4.49
Cr Cd ._*_'_
1.07 0.61 4.80
1.08 0.62 4.72
1.11 0.64 4.93
NOTE: 'Organic phase: 40 v/o LIX 64N. 60 v/o 470B contacted with leach solution (892) to pre-condition;
then stripped with synthetic electrolyte solution (30 gpl Cu. 180 gpl H.SOJ.
•All contacts performed for 3 minutes at 20°C.
-------
TABLE 8.70. CONTINUOUS COPPEK EXTRACTION FROM MIXED METAL LEACH SOLUTION BY
LIX 64N
Organic Phase: 10 v/o LIX-64N; 90 V/0 KERMAC 470-B
Aqueous Phase Composition (gpl): Cu Fe Cr Ni Zn ptl
2780 7703 0.64 6.86 8.13 Adjusted
to 2
Loading Contacts 3, 0/A = 1
Stripping 2. 0/A = 10
Flow Rate 100 cc/min.
Volume Treated 10 liters
Copper Content of the Raffinate 11 ppm
Copper Content of the Strip Solution 4.96 gpl
Acid in Strip 20 v/o H
used to develop the isotherm data followed that prescribed by Henkel
Corporation (30): The organic solution is loaded with a little copper by
contacting it with feed in a separatory funnel. Next, the organic is shaken
with a typical tankhouse electrolyte (30 g/1 Cu, 150 g/1 H^SO^) at an C/A = 1.
This pre-prepared organic is then contacted with aqueous leach solution at
various 0/A ratios. The two phases are recovered and analyzed.
8.6.1.2. LIX-622
LIX-622 is a reagent developed by Henkle Corporation for copper extraction
applications under acid conditions where the pH <1.5 and where the copper
content is rather high, i.e., and >2 gpl. The design matrix results (Table
8.71. and 8.72.) verify that effective copper extraction occurs within the pH
range 1.0-1.5. The effect of pH (in the range 1-1.5) is not very important
with respect to copper extraction, but it is important to keep the pH low in
order to minimize the extraction of other elements (which does occur to a
greater extent as pH is raised). From the effects portion of the design matrix
table it appears that pH is the only important variable influencing extraction
of elements other than copper.
302
-------
TABLE 8.71. DESIGN MATRIX FOR LIX-622 EXTRACTION OF COPPER FROM SLUDGE LEACil SOLUTION (1/2 REPLICA)
s
1
3«
315
316
30
318
im
350
351
)*)•!
Bill
Unit
High (.)
lo. (.)
Uil lo.
1
2
1
<.
5
5? 6
1
8
B»i
Effects
Cu
Fe
Cr
Ni
Zn
Cd
III 622
It)
15
5
20
iO
-
4
-
4
-
4
-
4
0.2
-0.1
-0.4
-0.1
-0.2
-1.0
lypi
(irotini
470-B
—
470-B
450
-
-
4
4
-
-
4
4
-4 5
-1.1
d A_
- 0
- .0
Oiconul
(I)
10
10
20
0
-
-
-
-
4
4
4
4
-14.0
0.9
- 1.1
- 1.1
- o a
- 1.0
MM
3
2
5
1
-
4
4
-
4
-
-
4
4.8
-0.4
-0.4
-0.4
-0.5
-10
Mil
l«lp.
40
15
55
25
-
-
4
4
4
4
.
-
-0.2
-IB
-2.4
-2.1
-IB
-1.0
llMh
SolutUn
pH
1.25
0.25
1.5
1.0
-
4
-
4
4
>
4
.
0.5
96
9.4
46
9 5
q 5
Results: Extraction from Solution (%)
•
Nil. lip.
Cu
88.0
97.0
86.0
80.0
70.0
58.0
50.0
61. 0
8). 0-4.0
i4.6
Fe
15.0
33.0
B.O
28.0
28.0
9.0(10.0)
30.0
10.0
IJ. 0-1.0
18.3
Cr
14.0
31.0
S.O
25.0
25.0
6.0(6.0)
28.0
8.0
10.0-2.0
i 6.5
V«riitloo
NOTE: -Sludge Type 2
•Initial Solution
0.66 Cu. 3.18 Fe.
2.29 In. 0.12 Cd
•Oraanic/Aaueous •
Ni
11.0
29.0
4.0
24.0
23.0
t.O(t.O)
26.0
6.0
J.Oil.O
sl.O
Zn
15.0
32.0
B.O
28.0
28.0
9.0(9.0)
30 0
10.0
12.0-2.0
16.0
Cd
8.0
25.0
0.0
17.0
17.0
o.n(o.o)
25.0
0.0
0.010.J
16.5
Composition (gpl):
0.36 Cr. 1.37 Ni.
1 : SOcc each
•Test 6 duilteated
•Baseline run three times
-------
TABLE 8.72 OBSERVATIONS ON PHASE SEPARATION: DESIGN MATRIX TESTS (TABLE 8.54)
FOR COPPER REMOVAL USING LIX 622
Test I
I
2
3
4
5
6
6b
• 7
B
Baseline A
Baseline B
Baseline C .
•Muck: A layer of organic-aqueous that disappears slowly.
Observations
Good Separation
Fair Separation, Mucky*
Good Separation
. Little Muck
. But Some Muck
-------
The design matrix study approach should only be considered a qualitative
evaluation of system experimental variables. The interpretation should be
limited to pointing out parameters that have a large effect on element
recovery. The design matrix results show reasonably high metal value
extraction of associated elements. This result, however, is due to the fact
that the tests were conducted prior to establishing the extraction ability of
the LIX-622 reagent, i.e., the extraction ability of LIX-622 is approximately
0.3 gpl Cu / v/o LIX-622 (see Table 8.73). Therefore, for the LIX-622 contents
used fn the design matrix a very large excess of reagent was present, e.g.,
even for the lowest concentration, 5 v/o LIX-622. the reagent has the ability
to pick up approximately 1.5 gpl Cu from the aqueous solution (the starting
solution contained only 0.66 gpl Cu). The data are Important, however, to
consider because they illustrate that even IP the presence of a large excess of
reagent that impurity pick-up by the organic Is controllable, e.g., for the
baseline condition lowering the pH of the aqueous phase from 1.25 to 1.0 should
decrease the extraction of copper by only 0.5% but should decrease the
extraction of all associated elements by 10% which would mean essentially no
associated elements would be extracted (even in the presence of a large excess
of reagent).
The above interpretation is confirmed in Table 8.74., where a high copper
bearing solution (7.54 gpl Cu) is contacted with a 25 v/o LIX-622 organic
phase, i.e., note that essentially no associated element is removed. The
concentration of LIX-622 should have been slightly greater to more effectively
remove all the copper.
The Influence of pH on copper extraction from a sludge leach solution
containing a nigh concentration of copper and iron is presented in Table 8.75.
Selective copper recovery from a mixed metal solution can be achieved from
iron bearing solution (before jarosite) or from iron free solutions (after
jaroslte). However, the phase separations are not as fast or as clean for the
before jarosite leach solutions.
305
-------
TABLE 8.73. COPPER LOADING SUMMARY IN LIX 522 (25 v/o) - KERMAC 4708
B
Sample No.
1044
1045
1046
1050
1047
1048
1051
1049
gpl Cu / v/o LIX 622 9 pll « 1.75
0/A
10 5 2 1 0.5 0.2
0.27
0.30
0.30
0.30
0.29
0.21
0.20
0.11
NOTE: -See Table 8.57 for nrocedure.
•Maximum copper loading In organic phase: 14.4 gpl.
-------
TABLE 8.74. LIX 622 ISOTHERM DATA: 25 v/o LIX 622 APPLIED TO JAROSITC TREATED SOLUTION
1
Sample No. Condition
1042
1043
1044
1046
10SO
1047
1048
1051
1049
Starting Solution
(Barrel Z sludge)
Jarosite Solution
UnfiUered leach 1042
subjected to potassium
Jarosite conditions for
4 hours. Final pH = 1.75
Isotherm (25 v/o LIX 622.
75 v/0 KERI1AC 470B "
Organic exposed to
solution 1043:
0/A - 10
0/A • 2
0/A • 2 (repeat)
0/A - 1
0/A - 0.5
0/A - 0.5 (repeat)
0/A • .2
Cu
7.54
7.69
0.10
0.17
0.23
0.53
2.33
2.68
4.81
Fe
22.08
6.53
6.56
6.71
6.SB
6.68
6.68
6.56
6.50
Concentration (gpl)
Zn
14.90
15.43
15.56
16.01
15.62
15.93
16.56
15.61
15.30
Hi
6.17
6.60
6.64
6.87
0.67
6.87
6.99
6.66
6.53
Cr
1.46
1.24
1.25
1.28
1.26
1.28
1.31
1.25
1.23
Cd
0.86
0.94
0.95
0.97
0.95
0.97
1.00
0.94
0.92
Al
5.69
4.72
4.77
4.B6
4.79
4.88
5.02
4.7P
4.70
-------
TABLE 8.74. CONTINUED
NOTE: -Standard leach on Barrel 2 sludge; 1/2 hr.. 50°C. pll • 1-1.5.
•Jarosite conditions: leach solutions plus leach solids subjected to conditions
of: 90°C. 4 hrs., K2S04/Fe • 1, Initial pH - 2.5.
•Organic phase: 2* v/o LIX 622, 75 v/o 470B contacted with 10 gpl Cu In aimonlca!
solution for 3 mln., then stripped with 180 gpl IUSO..
•All contacts performed In 125 cc separatory vessels fnr 3 minutes at 20°C.
-------
TABLE 8.75. INFLUENCE OF Pll ON COPPER EXTRACTION FROM SLUDGE lr*CH SOLUTION: LIX 622
Solution
ph( Initial)
1.00
(Feed-No Contact)
1.00
1.25
1.51
1.75
2.00
9P1
3.41
0.94
0.31
0.16
0.11
0.08
Concentration
Cu
Extracted
72.4
90.0
95.3
96.8
97.7
Fe
7.31
7.36
7.36
7.J7
7.25
7.16
in Solution After Contact (gpl)
Ni
2.66
2.66
2.68
2.69
2.73
2.69
Zn
5.20
5.23
5.24
5.20
5..*3
5.12
Cr
0.50
0.50
0.51
0.51
0.51
0.4?
Cd
0.37
0.38
0.36
0.38
0.39
0.37
NOTE: <10 v/o LIX-622
•90 v/0 KERMAC 470B
•0/A « 1, 60 cc each, concentration corrected for dilution by pH adjustment
•One contact
-------
8.6.2. Larg_e Scale Copper Extraction Test work
A large scale leach (1/10 design scale) was performed to sjpply leach
solution (14 liters) for the new one-gallon mixer solvent extraction system.
LIX-622 was used as the extractant. The results are presented in Table-8.76.
Copper was removed from the leach solution to a level of 43 mg/liter. ?ther
metal values were not extracted.
Results of large-scale testwork in the Reister SX system is presented in
Table 8.77. Copper can be selectively and effectively removed from the leach
solution. Phase separation is excellent and only after many hours of operation
using the same reagent does a muck l».yer build up at the interface. The muck
, layer can be easily withdrawn by simple aspiration. The muck layer is fine
particulate jaroslte carried into the system in the leach solution.
The large scale testwork was probably not run sufficiently long to
establish steady state conditions in the system. Henkel in their test work,
even in a small continuous system, states' ' that they run tests for periods
of at least 4-5 days to determine steady state conditions and to really
understand what element distribution 1s occurring during the transfer and to
understand if crud formation is going to be a problem. Long-term tests were
performed during the Phase II study.
Several associated studies were performed to understand more about the
control of the large scale SX system. One such study was to run the system
near 500 cc/min. to determine if the interface levels could be established and
maintained. The system was controllable and once the Interface levels were
established only minimum attention to adjustment was required. Chemical
results from the control test are presented in Table 8.78.
At one point in the study it was felt that aging time of the leach
solution may influence greatly the formation of crud and muck in the SX system.
This supposition was proved unfounded (at least for aging times of one month).
However, interesting small scale batch testwork was performed on aged solutions
for copper extraction by LIX-622. Studies were conducted on leach solution
(high iron) and jaroslte treatment (low iron) leach solution. The influence of
310 •
-------
TABLE 8.76. FIRST LARGE SYSTEM (ONE GALLON MIXER-SETTLER) 1EST FOR COPPER EXTRACTION USING LIX 622
Conditions: IS v/o LIX 622
R5 v/o KERKAC 470B
Two Stages of Extraction
One Stage of Strip
(4i of Leach Solution into Systen:
Temperature :
Solution Flow Rate :
Total Volume Treated :
Strip Acid :
Cu Fe Hi Zn Cr
Original Feed (gpl) 2.73 6.10 1.90 4.04 0.42
Rafflnate (gpl) 0.043 6.14 1.94 4.12 0.42
K25
25°C
2SO cc/ninute
14 liters
200 gpl H2S04
Cd '
0.24
0.25
-------
TABLE 8.77. COPPER EXTRACTION RY UX-622 DURING LARGE SCALE TESTHORK IN THE REISTER TESTRACK
Sample
1523
1524
1802
1797
1805
1816
2127
2129
2142
2144
2494
2499
Condition
40 Liter Test (10 v/o LlX-622)
Starling Solution. pH • 2.14
Raffii'iate (composite),
pH * 1.73
60 Liter Test (10 v/o LIX-622)
Starting Solution, \M • 1.9
Raffinate. 1/2 hour
" 2 hours
" 4 hours
90 Liter Test (10 v/o LIX-622)
Starting Solution, pH - 2.01
Raffinate. 2 hours (pH • 1.33)
* 6 hours
" Composite
160 Liter Test (15 v/o LIX-622)
Starting Solution, pH • 1.9
Raffinate (composite) pH • 1.3
Concentration (gpl)
Cu
1.37
0.017
0.39
0.007
0.065
0.022
3.89
0.38
1.34
0.78
3.05
0.030
Fe
0.65
0.6R
1.13
1.12
1.14
1.14
.
0.33
0.31
0.30
0.31
0.57
0.52
In
4.44
5.14
8.89
8.89
8.89
8.34
5.80
5.63
5.42
5.58
•
6.58
6.47
_Cr
0.28
0.29
0.26
0.26
0.26
0.27
0.36
0.35
0.33
0.34
3.08
2.95
Hi
3.02
3.18
8.02
7.87
8.00
8.07
3.15
3.06
3.07
3.20
1.67
1.68
Cd
0.31
0.33
0.41
0.41
0.41
0.42
0.44
0.42
0.42
0.44
0.11
0.11
Al
0.55
0.56
•
0 44
0.45
0.45
0.45
1.01
0.98
0.93
0.96
1.30
1.25
Continued
-------
TABLE 8.77. CONTINUED
NOTES: -Solutions treated for Iron removal by jaroslle precipitate prior to SX exposure.
•40 liter test conditions: 2-state extraction (C/A • 1)
1-stage strip (0/A • I. ISO gpl 117504)
Flowrate: 250 cc/rain. all phases
Temperature: 30-SO°C
•60 liter test conditions: 2-stage extraction (0/A • 1)
l-stage scrub (100 gpl KjSOd.)
1-stage strip (0/A • I, 175 gp
Flowrate: 250 cc/nin. all phases
Temperature: 30-5fl°C
•90 liter test conditions: 2-stage extraction (0/A • I)
2-stage strip (0/A « 1. 150 gpl H:S04
Flowrate: 250 cc/mtn. all phases
Temperature: 30-50°C
•160 liter test conditions: 2-stage extraction (0/A • 1)
2-stage sirip (0/A - I, 175 gpl H2S04)
Flowrate: 2t>0 cc/nin. all phases
Temperature: 30-SO°C
-------
TABLE 8.78- CHEMICAL RESULTS
Simple No.
Condition
ON LARGE SCALE
Cu SX CONTROL TEST
Concentration
_Cu__
2093
2092
2094
2095
Feed (mixture of 1466
and 1991)
Composite Rafflnate
(8 gallons - flow
rate 250 cc/mln.)
Composite Rafflnate
(7 gallons - flowrate
475 tc.'mln. -design
limit for system'
Strip from recycled
above two tests
3.
94
0.26
(NOTE:
0.070
(NOTE:
17.
22
Fe
1.58
1.23
System at
solution
Hi
6.48
Cr
0.59
(qpD
Zn
10.10
5.44 0.53 8.74
start-up contained 2 gal
.'. can't compare decrease
Cd
0.69
1
Al
.40
0.58 1.50
of previous leach
to 2093 feed)
1.57 6.43 0.59 10.05 0.68 1
System switched from 250 cc/mln. to 475 cc/mln.
solution composition In cells same as feed 2093
0.02
0.03
0.01
0.10
< D. L.
0
.40
so
.03
NOTE: -10 v/o LU 622, 90 v/o KERHAC 470B.
•Two stages of extraction (0/A • 1). two stages of strip (0/A - 1).
•System mixers each had a flowmeter Installed and flows controlled at designed rates.
-------
aging and solution dilution on LIX-622 extraction of copper was observed on
each type of solution. The results 'or the non-jarosited solutions are
presented in Table 8.79. The aging effect was very pronounced on undiluted
leach, i.e., less copper was extracted from the longer aged solutions. There
was not much difference between a five hour age (0.7 gpl Cu removed) and a 22
hour age (0.78 gpl Cu removed) but there was much more copper extracted from a
1.5 hour age (1.16 gpl Cu removed). The same trend was true of a 20 v/o
diluted solution but essentially no aging effect was noted when the solution
was diluted by 100 v/o.
The results for the jarosite treated leach solutions are presented in
Table 8.80. The aging effect on undiluted leach solution showed 20 percent
less pick-up of copper by a solution aged for 13 hours. There appears to be
essentially no aging effect for the 20 v/o dilution and 100 v/o dilution test
results.
Insufficient test work has been performed to establish the reason for the
apparent aging effect in some cases and not in others. The effect may be of
only academic interest because effective recovery has been found in the large
scale continuous test work (at least for jarosite treated solutions) for
solutions run without regard to storage time. Storage times ranged from a few
hours to a few days.
8.6.3. Long Term Copper Extraction Testwcrk
A series of studies were conducted to investigate the stage and process
efficiency and the possible degradation of the LIX-6Z2 bearing organic phase.
The system was described previously in Section 5.2.1. It consisted of three
stages of extraction and two stages of sulfuric acid stripping.
The tests were conducted in the Bell Engineering solvent extraction
testrack; 3.88 liters of 15 volume percent LIX-622 - 85 volume percent KERMAC
470B was contacted with 341 liters of aqueous leach solution over a period of
113 hours. Approximately 226 load/strip cycles were achieved. An
aqueous/organic contact ratio of over 88 was achieved. The results of the
315
-------
TABLE 8.79. INFLUENCE OF AGING TINE AND DILUTION
Sample No.
2054
2055
2058
u>
en
2061
6064
2067
2056
2059
2062
2065
Condition
Starting Leach Solution:
Undiluted, pn • 1.5
Repeat Analysis of 2054
Leach Aged 1.5 hrs.,
then Contacted with
LIX 622
Leach Aged 3 hrs., then
Contacted wTHTTlX 622
Leach Aged 5 hrs.. then
Contacted with LIX 622
Leach Aged 21 hrs.. then
Contacted with LIX 622
Starting Leach Solution:
20 v/o Dilution, pll = 1.5
Leach Aged 1.5 hrs., then
Contacted wTtnTTC 622
Leach Aged 3 hrs., then
Contacted wTTTTTlX 622
Leach Aged 5 hrs.. then
ON LIX
622 EXTRACTION FROM
Concentration
Cu
1.25
1.24
0.085
U.320
0.531
0.464
0.997
0.1 SO
OJJ8
0.319
Fe
12.01
11.84
11.67
12.21
12.25
12.42
9.43
9.70
9.74
10.00
H|
5.19
5.13
5.10
5.32
5.31
5.43
4.11
4.24
4.24
4.35
Zn
5.69
5.62
5.58
5.78
5.80
5.93
4.47
4.61
4.60
4.72
Cr
0.66
0.65
0.65
0.67
0.67
0.69
0.52
0.53
0.54
0.55
(gpn
Cd
0.40
0.40
0.39
0.41
0.41
0.42
0.31
0.32
0.32
0.33
Al
1.85
1.84
1.82
1.88
1.88
1.91
1.48
1.51
1.50
1.52
LEACH SOLUTION
SI
1.37
1.36
1.34
1.39
1.40
1.42
1.09
1.10
1.10
1.12
Ca
0.
0.
0.
0.
0.
0.
0.
0.
0.
60
59
58
61
61
62
48
49
49
51
Contacted wTEKTlX 622
-------
TABLE 8.79. CONTINUED
CJ
Sample No.
2091
2057
2060
2053
2066
2069
Condition
Leach Aged 22 hrs.. then
Contacted wTUTLTX 622
Starting Leach Solution:
100 v/o Dilution
Leach Aged 1.5 hrs.. then
Contacted with LIX 622
leach Aged 3 hrs.. then
Contacted wTltiTlX 622
Leach Aged 5 hrs., then
Contacted wTOTTlX 622
Leach Aged 22 hrs.. then
Contacted wTDTTTx 622
Concentration (gpl)
Cu Fe N1 Zn Cr Cd Al St Ca
0.274 9.89 4.3S 4.71 0.55 0.33 1.51 1.11 0.50
0.613 6.03 2.64 2.82 0.34 0.20 0.93 0.6C 0.33
0.054 5.87 2.55 2.78 0.33 0.19 0.91 O.C3 0.31
0.080 6.00 2.59 2.83 0.33 0.19 0.92 0.64 0.3?
0.053 5.80 2.52 2.73 0.32 0.19 0.89 0.62 0.31
0.067 6.05 2.65 2.86 0.34 0.20 0.92 0.65 0.32
NOTE: -Standard leach on barrel 18 material. Aging done in contact with leach solids.
•All Contacts made with 10 v/0 LIX 622. 90 v/o KERMAC 470B.O/A « 1. i - 25°C, 3 minutes.
-------
TABLE 8.80. INFLUENCE OF AGING TIME AND DILUTION ON LIX 632 EXTRACTION FROH JAROSITE LEACH SOLUTION
Sample No.
2077
2080
2083
2086
2078
2081
2084
2087
2079
2085
2088
Condition
Concentration (gpl)
Starting Leach Solution:
Undiluted , pH » 1.8
Leach Aged 1 hr., then
Contacted WTUTLIX 622
3 hrs.
13 hrs.
Starting Leach; 20 v/o
miution. oH^Ta
Learn Aged 1 hr., then
contacted wTTTTLIX 622
3 hrs.
13 hrs
Starting Leach: 100 v/o
Dilution, on . I.B
Leach Aged 3 hrs.. then
Contacted wTDTTlX 622
13 hrs.
Cu Fe Nl Zn
Cr Cd
Si Ca
0.71 0.68 5.55 6.50 0.35 0.45 1.41 1.53 0.65
0.094 0.75 5.54 6.47 0.35 0.45 1.41 1.53 0.65
0.086 0.67 5.41 6.47 0.34 0.44 1.38 1.50 0.62
0.117 0.68 5.56 6.49 0.35 0.45 1.41 1.53 0.61
0.562 0.58 4.38 5.13 0.28 0.35 1.11 1.20 0.53
0.049 0.61 4.33 5.11 0.28 0.35 1.10 1.18 0.52
0.042 0.58 4.36 5.09 0.28 0.35 1.10 1.17 0.52
0.064 0.59 4.43 5.14 0.28 0.36 1.12 1.20 0.53
0.346 0.36 2.69 3.15 0.18 0.21 0.68 0.70 0.34
0.018 0.37 2.79 3.26 0.18 0.22 O./l 0.74 0.34
i
0.016 0.38 2.89 3.34 0.19 0.23 0.72 0.75 0.36
NOTE: -standard Teacn onoarreT la maTcriai.
•Jarosite conditions: 6 hrs, initial pH > 2.2 . temp. • 90°C,
•Aging done in contact with solids.
•All contacts made with 10 v/o IIX 622. 90 v/o KERMAC 4708, 0/A • 1. T • 25°C, 3 minutes.
°
-------
study are summarized in Table 3.81: stage efficiency and process rack
efficiency in Table 8.82 and conditions for the testwork in Table 8.83.
Complete experimental results are presented in Table 8.84. The results of
degradation testwork performed on organic samples were presented previously in
Tables 6.19 and 6.20.
Copper is effectively and selectively entrailed from a mixed metal
solution (Table 8.81) by an organic phase that has been exposed to a large
number of load/strip cycles. Degradation does not appear to be a problem over
the test period studied. Stage efficiency for copper extraction from the
aqueous phase (Table 8.82} decreases with number of stages. This is expected
because the pH of the aqueous phase decreases as it moves from one stage to the
next. The overall process efficiency for copper extraction Is excellent and
low copper bearing solutions are produced (Table 8.82).
The physical separation of phases in the settlers is rapid and without
muck problems. A small amount of crud (sollas) forms but it remains Intact at
the first Interface and its source is most likely fine particulate solids
carried over from the leach solid/liquid separation unit operation. The crud
Is easily removed by aspiration from the Interface.
8.7. ZINC SOLVENT EXTRACTION (HIGH IRON FLOWSHEET)
A discussion was presented previously for a flowsheet to treat iron
bearing solutions (a few grams per liter): Section 8.4. The discussion In this
section relates to the nigh iron sludge treatment flowsheet that includes
removal of most of the iron by a jarosite precipitation process; removal of
copper by LIX-622 solvent extraction; then removal of zinc and residual Iron (a
few hundred mg/l). Therefore, the discussion of zinc extraction by SX Is
limited in this section to treatment of jarosite treated leach solutions.
A design matrix test series has been performed on a copper and iron free
solution. The extraction results are presented in Table 8.85. (phase
separation notes in Table 8.86). The solution used in this test series was
exceptionally low in zinc content (1.24 gpl). Typical zinc contents of leach
solutions are in the range of 3-5 gpl.
319 .
-------
TABLE 8.81. LONG TEKM DEGRADATION
Sample No.
Condition
Organic Aqueous Mixer
Exposure Time.
3459
3475
3482
3493
3501-8
3509
3519
3538-A
3542
3547
*
46.5
46.5
40.0
86.5
39.0
125.5
36.0
161.5
25.5
187.0
Hrs. Feed
First Day
Starting Feed
Solution
15.5 279
Second Day
Starting Feed
Solution
28.8 518.4
Third Day
Starting Feed
Solution
41.8 752.4
Fourth Day
Starting Feed
Solution
53.8 968.4
Fifth Day
Starting Feed
Solution
62.8 1121.4
• EXTRACTION
STUDY:
COPPER BY 15 V/0 LIX 622
Concentration In Final Composite RaFflnate. gpl
Contact
(mln.)
Strip
2.
186 0.
3.
345.6 U.
3.
501.6 0.
3.
645.6 0.
2.
747.6 0.
Cu
750
054
130
062
130
106
332
088
600
039
Fe
3.899
3.727
4.068
3.995
4.068
3.958
3.809
3.918
4.237
4.299
Zn.
0.111
0.131
0.129
0.131
0.122
0.127
0.097
0.130
0.148
Cr
1.987
1.943
2.081
2.086
2.084
2.060
2.030
2.067
2.238
2.224
Ni
5.147
5.750
6.260
6.150
6.260
6.291
5.749
5.951
6.815
6.705
Al
0.207
0.217
0.248
0.259
0.248
0.281
0.353
0.387
Ca
0.317
0.318
0.331
0.307
0.331
0.344
0.369
0.335
P
0.572
0.570
0.656
0.534
0.656
0.607
0.652
0.689
0.627
0.633
-------
TABLE 8.81,
Sample KD. Condition
Organic Aqueous Mixer Contact
Exposure Time, (nin.) Cu
3552
3567
3606
3613
3619
3631
3639
3643
3651
3664
t Hrs. Feed Strip
Sixth Day
19.5 Starting Feed
Solution
206.5 68.8 1238.4 825.6
Seventh Day
34.6 Starting Feed
Solution
241.0 79.3 1427.4 951.6
Eighth Day
34.5 Starting Feed
Solution
275.5 90.8 1634.4 1089.6
Nintn Day
12.0 Starting Feed
Solution
287.5 94.3 1706.4 1137.6
Tenth Day
27.0 Starting Feed
Solution
314.5 103.8 1868.4 1245.6
0.835
0.056
1.035
0.033
2.045
0.027
1.812
0.073
2.026
0.043
, CONTINUED
Concentration In Final Composite Rafflnate. gpl
Fe
3.471
3.373
3.616
3.546
3.117
3.069
•3.185
3.125
2.746
3.609
Zn
0.118
0.091
0.126
0.105
0.099
0.096
0.128
0.121
0.079
0.122
tr
1.813
1.788
1.907
1.940
1.725
1.725
1.739
1.959
1.515
1.879
Hi
5.417
5.251
5.727
5.574
5.019
5.002
5.100
5.652
4.464
5.866
Al
...
0.362
0.405
0.354
0.355
0.361
0.406
0.318
0.386
Ca P
0.279
0.288
0.322
0.351
0.303
0.284
0.316
0.348
0.356
0.330
-------
Sample
No. Condition
TABLE 8.81
. CONTINUED
Concentration In Final Composite Rafffnate, gpl
Orjinic Aqueous Mixer Contact
Exposure Time, (mtn.) Cu
3670
3703
t Mrs. Feed
Eleventh Day
Strip .
27.0 Starting Feed 2.225
Solution
341.5 112.8 2030.4 1353.6 0.049
(226 Load/Strip Cycles) ~
Fe
3.078
3.126
Zn Cr Ml At Ca P
0.105 1.693 5.046 0.37J 0.276
0.112 1.687 5.071 0.383 0.279
NOTES: Conditions given for each day s testwork in Table 8.83.
'Total organic in extraction system « 2338 cc; In strip system 1550 cc.
'Total aqueous In extraction system • 2055 cc; in strip system 1370 cc.
-------
TABLE 8.82. CELL EFFICIENCY FOR COPPER EXTRACTION BY LIX 622: LONG TERN TEST (Lew Iron Flowsheet)
Sample Ho. Conditions Stage Efficiency, gpl in Rafflnatc. I Extracted
Cell 1 Cell 2 Cell 3 Efficiency
gpl t gpl * gpl *
3459 Starting Solution 2.750 gpl
(pH • 1.75)
3463 6.0 Hrs. Continuous Exposure 0.216 92.1
3464 " • ' ' 0.032 85.2
3465 ' ' ' • 0.018 43.8 99.3
346? 13.5 Hrs. Continuous Exposure 0.345 87.4
3470 ... . 0.038 89.0
3471 ... • o.022 42.1 99.2
3482 Starting Solution 3.130 gpl
(pll - 1.75)
3486 21.5 Hrs. Continuous Exposure 0.867 72.0
3487 .... 0.1H fl2.4
3485 ... . o.053 65.4 98.3
3501-B Starting Solution 2.697 ,
(pH - 1.76)
3504 34.8 Hrs. Continuous Exposure 0.483 82.1
3505 ... . 0|Q68 35.9
3503 ... . o.072 97.3
3519 Starting Solution 2.700
(pH • 1.75)
3526 47.8 Hrs. Continuous Exposure 0.734 72,8
3527 ... . o.094 87.2
3530 ... . 0.097 96.4
-------
S3
TABLE 8.82. CONTINUED
Sample No.
3542
3544
4545
3546
3606
3608
3609-A
3610-A
3619
3632
3S33
3631
3551-A
3658
3659
3660-B
Conditions Stage Efficiency, gpl In Raffinate,
Cell 1 Cell 2 Cell
gpl * gpj^ __»_ gpl
Starting Solution 2.600
(pH • 1.77)
57.8 Hrs. Continuous Exposure 0.233 83.8
' " " ' 0.040 82.8
' ' " ' 0.040
Starting Solution 1.035
tpll - 2.0)
73.8 Hrs. Continuous Exposure 0.027 97.4
• " ' ' 0.036
" • ' " 0.026
Starting Solution 2.045
(pll • 2.0)
84.3 Hrs. Continuous Exposure 0.142 93.1
• ' " " 0.031 78.2
• ' 0.027
Starting Solution 2.026,
(pH - 2.1)
99.8 Hrs. Continuous Exposure 0.593 76.7
• ' ' ' 0.053 91.1
• • • - — o.ois
t Extracted
Total
3 Efficiency
1
0 98.5
27.8 97.5
IZ£ 98.7
•
71.7 99.4
3670 Starting Solution 2.225 gpl,
(pH • 2.0)
-------
TABLE 8.82. CONTINUED
Sample Mo. Ccntfltlons Stage Efficiency, gpl In Baffinate. 1 Extracted
Total
Cell I Cell 2 Cell 3 Efficiency
gpl _ t gpl * gpl %
3700 112.8 Hrs. Continuous Exposure 1.116 49.8
3701 ... . 0<266 76.2
3702 ... . 0 068 77.4 96.9
M
in ^~~——^——~--~-~-•"—^~~~•—-~~-•""•—~-~-~-~~———•—~~~~-~~~~~~~•"~~—~~~^~^~~~•
NOTE: Detailed data presented tn Table 8.81.
-------
TABLE 8.83. CONDITIONS FOR TABLE 8.82 TESTWORK
o>
Identifier Test Numbers
3459,
3482.
3474
3493
3501 -B. 3509
3519.
3542.
3552.
3606.
3619.
3b39,
3651.
3670.
3533
3548
3567
3613
3631
3643
3664
3703
Description
Initial
Cu. gpl
First Exposure of Organic 2.750
Second Exposure of Sane Organic 3.130
Third
Fourth
Fifth
Sixth
Seventh
Eighth
Ninth
Tenth
2.697
2.7CO
2.6CO
0.835
1.035
2.045
1.812
2.026
Eleventh 2.225
Conditions
Initial
pH. R.T.
1.75
.75
.75
.75
.77
.76
2.01
2.01
2.13
2.14
2.01
Volume
Average. 1
46.5
40.0
39.0
39.0
25.5
19.5
34.5
34.5
12.0
27.0
27.0
Temp.
Fe«d 50°C
Cell 1 35
Cell 2 30
Cell 3 27
Cell 4 24
Cell 5 24
Feed 28°C
Feed 45°C
Feed 44-55°C
Feed 50°C
Feed •
Feed 25°C
-------
tt
TABLE B. 84. EXPERIMENTAL RESULTS FOK LONG TERN ORGANIC EXPOSURE TESTHORK:
Sample No.
3459
3458
J460
3461
3462
3463
3464
3465
3466
3469
3470
3471
3472
3473
Conditions
First Pay (46 1 aqueous leach)
Starting Solution
Starting Strip Acid, 200 gpl
1.5 Hours
Cell 3
Composite
Cell 3
Cell 1
Cell 2
Cell 3
Composite
Rdffindte
3.0 Hours
6.0 Houis
7.5 Hours
Rdtflnate
13.5 Hours
Cell 1
Cell 2
Cell 3
Composite Rafflnate
Stria Acid
COPPER EXTRACTION WITH LIX 622
Concentration in Final Composite Raffinate. gpl
3
3
4
4
3
3
3
3
3
3
3
o
Fe
.899
.920
.024
.120
.994
.758
.675
.764
.899
.792
.859
.866
1114
Cu
2.750
0.027
O2T
0.025
0.216
OJ7
ore
0.081
0.345
OJff
O27
OzT
»a~r-
Zn
0.102
0.004
0.100
0.108
0.090
0.117
0.112
0.117
0.110
0.102
0.106
0.113
0.112
Cr
1.987
0.002
2.0i6
2.039
2.124
1.996
1.927
1.922
1.908
1.959
1.960
2.002
1.971
N<
5.847
5.903
6.033
6.135
5.895
5. 755
5.622
5.706
5.781
5.785
5.957
5.891
Al
0.207
0.209
0.218
0.224
0.243
0.230
0.220
0.214
0.223
0.230
0.240
0.225
Ca
0.317
0.326
0.130
0.332
0.378
0.315
0.3U4
0.310
0.318
0.3*9
0.322
0.321
P
0.572
0.660
0.694
0.696
0.688
0.604
0.625
0.611
0.580
0.602
0.631
0.677
3475
15.5 Hours
Final Composite
3.727 0.054 0.111 1.940 5.750 0.217 0.318 0.570
-------
TABLE 8.84.
Sample No. Conditions
348Z
3484
3486
3487
3488
3490
3491
3489
3493
3492
35U1-D
3504
3505
3503
Second Day (4u 1 aqueous leach)
Starting Solution
5.0 Hours
Composite Raffinate
6 Hours
Cell 1
Cell 2
Cell 3
13.1 Hours
Cell 1
Cell 2
Cell 3
Final Raffinate
Strip Acid
Third Day (39 1 aoueous leach)
Starting Solution
6 Hours
Cell 1
Cell 2
Cell 3
F«
4.068
4.007
3.877
3.932
3.901
3.961
4.180
4.010
3.995
0.052
4.033
3.993
3.916
3.640
CONTINUED
concentration In Final Composite
Cu
3.1_30
0.047
0.867
O75
1.218
OT?27
099
OS?
2.697
0.483
OEff
iHJT?
0.
0.
0.
0.
0.
oooo
....
0.
0.
0.
0.
2n
131
122
125
114
126
122
105
127
129
120
103
112
109
Cr
• ^^— ^^^
2.084
2.093
1.989
2.008
2.002
2.039
2.171
2.076
2.086
2.014
2.066
2.018
1.927
. "'
6.260
6.200
5.995
6.059
5.9b6
6.050
6.394
6.788
6.150
6.057
6.020
6.143
S.923
Raffinate. gpl
0.
0.
0.
0.
0.
0.
0.
0.
0.
Al
248
243
245
249
244
263
257
259
242
0.293
0.270
Ca
• ^BW^HH
0.331
0.316
0.316
0.301
0.301
0.302
0.308
0.321
0.307
0.324
0.347
0.353
0.334
p
0.656
0.389
0.760
0.602
0.660
0.616
0.645
0.647
0.534
0.700
0.648
0.623
0.607
-------
10
TABLE 8.84.
Sample No.
•
3506
3508
3510
3511
3509
3512
Cell
Cell
Conditions
7 Hours
3
10 Hours
3
13 Hours
Cell 1
Cell 2
Cell 3
Strip Acid
CONTINUED
Concentration in Final Composite Raffinate. gpl
Fe
3.351
3.965
3.925
3.904
3.958
0.052
Cu
•• ^MM«^
0.078
0.091
1.072
OTZDT
OTTO?
3OT
» ••
0.
0.
0.
0.
0.
Zn
114
111
130
120
122
Cr
• OI^H^^^^HM
1.986
2.044
2. 059
2.098
2.060
HI
A OBB^lBflHi^
6.043
6.113
6.095
6.154
6.291
Al
• m^^^^^f^^m
0.265
0.294
0.290
0.288
0.281
Ca
• Bm^^H^^M
0.335
0.347
0.361
0.358
0.344
P
0.605
I
0.619
0.716
0.639
0.607
Fourth Pay (36 1 aqueous leach)
3525
3526
3527
3530
3531
3532
3533
3538-A
Cell
Cell
Cell
Cell
Cell
Cell
Cell
Final
3 Hours
3
6 Hours
1
2
3
12 Hours
1
2
3
Couposlte
3.786
3.954
3.987
4.165
4.140
4.339
3.918
0.041
0.734
oir
O97
0.910
OUT
O.OUB
0.
0.
0.
0.
0.
0.
0.
112
120
115
119
131
127
097
2.029
2.106
2.093
2.209
2.127
2.258
2.067
6.042
6.273
6.342
6.450
6.494
6.969
5.951
0.283
0.299
0.309
0.331
0.305
0.309
...
•
0.319
0.323
0.331
0.378
0.370
0.394 •
...
0.667
0.807
0.771
0.779
0.777
0.713
0.689
-------
TAIJLE 8.84. CONTINUED
Sample No. £°Miy£!l$ Concentration in Final Composite Raffinale, gpl
Fe Cu 2n Cr Ml Al Ca P
Fifth Day {25 1 aqueous leach)
3542 Starting Solution 4.237 2.600 U.13U 2.238 6.81S 0.3S3 0.369 0.627
3 Hours
3S43 Cell 3 4.247 0.048 0.127 2.228 6.810 0.378 0.367 0.678
4 Hours
3544 Cell 1 4.133 0.233 0.130 2.190 6.589 0.356 0.362 0.678
3545 Cell 2 4.103 OTO 0.137 2.185 6.752 0.357 0.362 0.533
3546 Cell 3 ' 4.423 OTO 0.134 2.297 7.046 0.419 0.354 0.5C9
8.5 Hours
3547 Cell 3 4.299 0.039 0.148 2.224 6.705 0.387 0.335 0.633
3549 Strip Acid 0.078 417158"
Sixth Day (20 1 aqueous leach)
3552 Starting Solution 3.471 0.834 0.118 1.813 5.417 — 0.279
3 Hours
3553 Cell 3 3.404 0.034 0.131 1.831 5.552 0.354 0.289
6.b Hours
3557 Cell 3 3.373 0.056 0.091 1.788 5.251 — 0.208 0.616
-------
TABLE 8.84. CONTINUED
t Sample Ho. Conditions Concentration In Final Composite Raffinate. gpl
Fe Cu Zn Cr Hi Al Ca P
Sevenin Day (34 1 aqueous leach)
3606 Starting Solution j.616 1.035 0.126 1.907 5.727 0.362 0.322
3 Hours
3607 Cell 3 3.609 0.029 0.126 1.952 5.918 0.319 0.337
5 Hours
«•» 3608 Cell 1 3.835 0.027 0.131 2.070 6.113 0.406 0.351
3609-A Cell 2 3.681 O35 0.120 1.992 5.817 0.426 0.365
3610-A Cell 3 3.519 OZS 0.130 1.932 5.600 0.390 0.346
10.5 Hours
3610-8 Cell 1 4.038 0.058 0.129 2.190 6.346 0.480 0.385
3611 Cell 2 3.85' OTC 0.119 2.093 6.142 0.447 0.360
3613 Final Composite 3.546. Ojl 0.105 1.940 5.574 0.405 0.351
3614 Strip Acid 0.137 36^75"
Eighth Pay (34 1 aqueous leach)
3619 Starting Solution 3.117 2.0M 0.099 1.725 5.019 0.354 0.303
3 Hours
3620 Cell 3 3.067 0.039 0.114 1.717 4.982 0.361 0.314
6 Hours
3628 Cell 1 3.230 1.775 0.107 1.782 5.186 — 0.307
3629 Cell 2 3.152 0.030 0.101 1.731 5.003 0.351 0.296
-------
TABLE 8.84.
Sample Ho. Conditions
3627
36J2
3633
3631
co 3634
CO
INJ
3639
3640
3641
3642
3644
3645
2543
3646
Cell 3
11.5 Hours
Cell 1
Cell 2
Cell 3
Strip Acid
ninth Day (34 1 aqueous leach)
Starting Solution
4 Hours
Cell 1
Cell 2
Cell 3
b Hours
Cell 1
Cell 2
Cell 3
Strip Acid
Tenth Day (27 1 aqueous leach)
Fe
3.193
3.334
3.165
3.069
0.146
3.185
3.494
3.477
3.443
3.640
3.670
3.525
0.319
CONTINUED
Concentration in Final Composite Raffinate. aol
Cu_.
0.036
0.142
O3T
O27
35757"
1.812
0.387
OJJ
on
1.179
OT36T
O7J
33TZB"
Zn
~ •••«•••
0.111
0.116
0.113
0.096
0.128
0.126
0.111
0.019
0.140
0.127
0.121
Cr
5.785
1.81:)
1.735
1.725
1.739
1.874
1.877
1.845
1.973
1.999
1.959
Ml
5.146
5.344
5.033
5.002
5.100
5.458
5.581
5.392
5.886
5.936
5.652
Al
0.356
0.389
0.359
0.355
0.361
0.398
0.399
0.381
0.443
0.409
0.406
Ca P
0.29?
0.319
0.294
0.284
0.316
0.345
0.337
0.335 —
• 0.347
0.345
0.348
3651 Starting Solution 2.746 2.026 0.079 1.515 4.464 0.318 0.356
-------
CJ
TABLE 8.84.
Sample No. Conditions
3653
3654
3652
3656
3657
36«
3662
3663
3669
3669
3670
3672
3673
3674
4 Hours
Cell 1
Cell 2
Cell 3
6 Hours
Cell 1
Cell 2
Cell 3
9 Hours
Cell 1
Cell 2
Cell 3
Strip Acid
Eleventh Day (27 1 aqueous leach)
Starting Solution
3 Hours
Cell 1
Cell 2
Cell 3
_fe_
3.553
3.196
3.398
3.396
3.553
3.305
3.691
3.682
3.609
0.269
3.078
3.035
3.133
2.956
CONTINUED
Concentration in Final Composite Rafflnate. gpl
Cu
*• m^^^m^m^^
0.197
on
02?
0.181
OZT
OJ5
0.954
O36"
O43
3OT
2.225
1.037
O47
Oft
• ^
U.
0.
0.
0.
0.
0.
0.
0.
6.
0.
0.
0.
0.
Zn
120
109
125
114
109
109
130
131
122
105
088
096
100
v i
1
1
I
1
1
1
1
1
1
1
1
1
1
Cr
•^to^^vi^^M
.880
.731
.853
.802
.884
.772
.971
.961
.879
.693
.642
.676
.638
Hi
H ^^H^^V^^^
5.644
5.006
5.406
5.379
5.560
5.336
5.721
5.883
5.866
5.046
4.869
5.006
4.859
Al
• «^^^^^^M
0.393
0.362
0.394
0.384
0.400
0.368
0.408
0.405
0.386
0.373
0.347
0.368
0.344
Ca
M ^^f^^f^m^mm
0.402
0.402
0.371
0.382
0.323
0.401
0.325
0.328
0.330
0.276
0.269
0.278
0.277
P
...
...
...
...
...
...
-------
TABLE B.84. CONTINUED
Sample No.
CiMltlons
Concentration in Final Composite RafMnate. gpl
3677
3678
3679
37UO
3701
3702
3703
Cell
Cell
Cell
Cell
Cell
Cell
Flna
6 Hours
1
2
3
9^ Hours
1
2
3
*i Composite
J
3
3
3
3
3
3
Fe
.024
.140
.017
.245
.064
.125
.126
Cu
1.116
O53
Of*
1.116
O5T
O5I
IT.U49
Zn
0.094
0.110
0.097
0.112
0.099
0.094
0.112
Cr
1.645
1.665
1.640
1.727
1.682
1.691
1.687
Ni
4.814
4.912
4.874
5.230
5.006
4.979
5.071
A1
0.36S
0.368
0.353
0.389
0.379
0.376
0.382
Ca
0.276
0.272
0.260
0.294
0.277
0.272
0.277
P
...
...
...
...
...
...
HOTES!'Test conditions In Table 8.81 and Table BTBTT
-------
to
TABLE 8.85.. DESIGN MATRIX FOR ZINC REMOVAL BY OEHPA FROM COPPER AND IRON FREE SOLUTION
s
•
i
ss?
SSI
SSI
sss
556 •)
we
559
se:
b&i-
Bin
Unit
HigS (.)
U. (-1
Ictl »
1
2
3
4
S
6
7
8
6t:<
Effects (5)
Zn
OlitP*
(»)
20
5
25
15
-
+
-
»
-
*
-
t
7.2
Otcenol
(X)
to
10
20
0
-
-
*
t
-
-
t
*
-19.7
ConUcl
lilt
(•in.)
3
2
5
1
•
-
-
-
i
t
*
«
-2.6
Conlici
lup.
(°)
40
IS
65
25
-
«
*
-
t
-
-
«
-0.6
Solution
»H
1.9S
0.2S
2 20
1.70
-
-
i
t
t
»
-
-
5.1
Results: Extraction from Solution (S)
.
Zn
68.5
81.8
36.5
57.4
IS.9OM]
86.2
26.0
37.0
80, MB?.)
NOTE: •
Cr
9,
7.
)6.
1
is. mi. n
IT. 4
11.0
-4.7
7.9 (?.])
Initial solution composition (gpl):
1.24 Zn, 0.43 Cr, 1.91 Nt. 0.19 Cd,
•Kertnac 470B make-up dllutent
•0/A « 1. 25 cc each
•Temp: 40°C
-------
TABLE 8.86. OBSERVATION ON PHASE SEPARATION FOR DEIIPA DESIGN MATRIX TEST (TABLE 8.64)
Tost I Observation
1 Very good phase separation, very little muck*
2 Very good phase separation, very little muck
3 Very good phase separation, very little muck
4 Very good phase separation, very little nick
5 Very good phase separation, little muck, very rapid
6 Very good phase separation, little muck, very rapid
ui
o> 7 Very good phase separation, little muck, very rapid
8 Very good phase separation, little muck
Base . Very good phase separation, no muck
*Huck • a layer of organic-aqueous that disappears slowly.
-------
A series of preliminary :hake tests to investigate zinc extraction from a
jarosite treated leach solution was conducted. Table 3.87. The results are
that zinc can be selectively extracted from a copper, iron free solution.
Phase separation is very good for iron free solutions, but were very poor in
the earlier work on zinc-iron solutions. The zinc extraction levels attained
In the shake test were not very low because there was insufficient 0_EHPA
present. Later experimental studies showed that about 0.15 gpl Zn is extracted
by each one volume percent D-EHPA; Table 8.88. Therefore, to totally remove
the zinc from a 7.45 gpl solution would require 50 v/o O.EHPA. Mote that the
data In Table 8.87. also shows that Deconol uecreases zinc extraction (in
agreement with the design matrix testwork reported in Table 8.85.). D^EHPA
Isotherm data are presented in Table 8.89. These tests were performed with an
organic phase deficit in sufficient quant-:>y of D-EHPA to completely remove the
zinc. It also appears that some N1, Cd, aid A! are partially extracted.
Extraction of Ni is not born out in large scale restwork. Aluminum and calcium
(not snown in Table 8.89.) are co-extracted.
The shake testwork was followed by experimental work in the Bell
Engineering 600cc continuous system. Results of a typical sequence test are
presented in Table 8.90. The continuous testwork was run under D^EHPA
deficient conditions, I.e., according to data generated later in the study
0.15-0.17 gpl Zn are extracted per v/o DgEHPA. Therefore, 40 v/o DgEHPA should
be able to maintain extraction of 6.0-6.8 gpl Zn. However, D^EHPA also
extracts ferric iron, some Al and Ca. Iron in the organic is not stripped so
in a recycle system extraction sites are occupied and are unavailable for zinc
extraction; also, Al that is extracted with the organic is only partially
stripped by 200 gpl H-SO^. Calcium loaded into the organic phase precipitates
when the organic phases cycles to the strip cells as gypsum. It can, however,
be easily filtered from the aqueous phase and removed from the system. The
data collected in the above testwork illustrates that long-term exposure of an
organic phase to a leach solution is needed to establish the required bleed
stream necessary to maintain an efficient regenerated 0-EHPA phase.
Large scale testwork in the Reister system supports this conclusion. The
results of four major tests are presented in Table 3.91.
•337
-------
TABLE 8.87. OEMPA Solvent Extraction Applied to a Jaroslte Treated Solution
Sample Condition J Concentration (gpl)
Fe Cu Cr Nl Zn Cd Al
Starting Solution
1300 Solution after potassium 0.33 4.51 0.38 2.07 6.9J 0.30 3.68
Jaroslte precipitation
IpH adjusted to 1.7S)
Copper Removal
Two contacts between 0.36 0.008 0.41 2.22 7.45 0.33 4.07
solution and LIX £22
(20 v/o). 0/A • 2, 3 niin.
DEHPA (401, 601 4708):
40°C
1 305
1 306
1 307
1 308
1 309
1310
I3ll
13I2
1313
1314
1315
0/A • 1 0.044
0/A • 2 0.036
0/A > 2 (repeat) 0.033
OEHPA (40X, lOt DEC,
50 v/o 470 B): 40°C
0/A • 1 0.048
0/A • 2 0.039
0/A • 2 0.033
OEHPA (40 v/o, 60 v/o
470 B). H702 Oxidized. 40°C
0/A • 1
-------
TABLE 8.88. SUMMARY OF ZINC LOADING FOR 40 v/0
Ul
3
10
PH • 2.0 0.19
pH • 2.C 0.17
pH • 3.0 0.18
DEIIPA. 60 v/o KERMAC 4708
gpl Zn/v/o OEHPA
5 2 _J 0.5 ^>.2
0.15 — - 0.10
0.17 0.17 0.15 0.13 0.09
0.12 -— 0.09
NOTE: 'Max. organic loading approximately IB gpl Zn.
•Organic solution pre-prepared by cont&ctlng with 100 gpl Zn solution, then stripping
with SO gpl Zn, 200 gpl H2S04 (0/A - 1). HH "
-------
TABLE 8.89. DEHPA ISOTHERM DATA-
40 v/o DEHPA. 60 v/o KERHAC 470B
Concentration (gpl)
1926
1928
1929
1930
192S
1918
1919
1920
PH - 2.0
Starting Solution
0/A • 10
Aqueous
Organic
0/A • 1
Aqueous
Organic
""" 0/A • 0.2
Aqueous
Organic
pH • 2.S
Starting Solution
0/A • 10
Aqueous
Organic
0/A » 5
Aqueo-js
Organic
0/A* 2
Aqueous
Organic
•- Zn
8.68
1.06
0.76
2.63
6.05
4.84
18.80
8.46
1.5?
0.69
1.47
1.40
1.78
3.34
Fe Cu
I. IS 0.002
0.71 <0. L.
0.04
0.9S
-------
TABLE 8.89. CONTINUED
I
1924
1921
1922
1923
1927
1931
1932
Concentration (gpl)
In
Fe
Cu
Ml
Cd
Al
0/A • 2 (Pepeat of 1920)
Aqueous
Organic
0/A • 1
Aqueous
Organic
0/A • 0.5
Aqueous
Organic
0/A • .2
Aqueous
Organic
pH • 3.0
Starting Solution
0/A • 10
Aqueous
Organic
0/A * 1
Aqueous
Organic
1.60
3.43
2.33
6.12
3.11
10.69
4.91
17.72
8.33
0.94
0.74
3.41
4.92
O.B5
0.14
1.02 •
0.11
0 93
0.40
0.99
0.70
1.12
0.67
0.04
0.93
0.09
<0. I.
• • • •
<0. I.
• • • •
<0. L.
....
<0. L.
....
0.002
<0. L.
....
<0. L.
8.2S
0.4S
10.39
8.34
1.60
9.01
0.6S
9.07
8.4S
0.06
8.31
0.76
0.36
0.06
O.S1
0.03
0.42
0.12
0.46
0.10
0.47
0.22
0.02
0.42
O.OS
0.29
0.06
0.33
0.09
0.35
0.14
0.38
O.?0
0.42
0.27
0.02
0.34
0.08
-------
TABLE 8.89. CONTINUED
Zn Fe Cu Ml Cd A)
1932 0/A • 1
Aqueous 3.41 0.93 < 0. L. 8.31 0.42 0.34
Organic 4.92 0.09 —- 0.76 0.05 0.08
1933 0/A • 0.2
Aqueous 4.70 1.00 < D. L. 9.24 0.4? 0.38
Organic 18.15 0.60
NOTE: 'Organic pre-prepared by contacting with 100 gpl Zn solution, then stripping (0/A • 1) with
50 gpl Zn. 200 gpl H.,S04.
uj •Temperature: 25°C.
ro •
-------
CJ
TABLE 8.90. OEHPA SOLVENT EXTRACTION OF ZINC: SHALL SCALE CONTINUOUS TEST
Sample
i
2005
2096
2097
1108
2109
2110
Conditions
Small SX System (600 cc mixer-
settlers)
Starting Solution
Initial pH « 2.0. aqueous
phase pH readjusted to 2.0
after two extraction contacts
Raffinate from Stage Two after
3 hours
Raffinate from Stage Four
after 3 hours
Raffinate from Stage Two
after 7 hours
Raffinate from Stage Four
after 7 hours
Strip after 7 hours of
recycling (initially
200 gpl H2S04>
Zn
5.7
OJL
0.029
L2L
0.15
33.05
Fe
•
0.26
0.08
0.03
0.10
0.06
-------
TABLE 8.91. ; ZINC EXTRACTION BY OCHPA DURING TESTWORK IN THE REISTER TESTRACK
Sample
1524
1532
1533
1811
1814
1124
1826
1827
1833
Condition
40 Liter Test (27 v/o DEHPA)
Starting Solution, pll -1.75
Rafflnate. pll • 1.29
Final Strip Solution
60 Liter Test (40 v/o OEIIPA)
Starting Solution, pH • 1.75
Raffinate, 1.5 hrs.
Raffinate. 3.25 hrs.
Above rafflnate adjusted to
pH « 2.0 and recycled
through system
pH Adjusted Feed
Rafflnate, 1 hour
2.75 hours
Concentration (gpl)
Zn
5.14
00
O4
8.84
177?
TST
1.61
07W
OTOT
Fe
0.68
0.46
0.005
1.13
0.97
1.00
0.97
0.60
0.60
Cu
0.017
C.014
2.5 and recycled through
system (strip acid replaced
with 200 gpl H2$04)
Feed
Rafflnate,
1 hour
3 hours
0.58
«OTCT
0.60
0.12
0.27
0.25
0.25
8.67
8.25
8.17
0.37
0.17
0.13
0.06
0.004
D.L.
Continued
-------
tn
TABLE 8.
Sample
-H^H>M__
2177
2178
2179
2181
21 BOB
2242
2246
2246
2253
2256
2255
2525
Condition
90 Liter Test (40 v/o DCHPA)
Starting Solution, pH • 2.0
Raffinate after second stage
contact (see note below)
Raffinate after fourth stage
contact
final composite raffinate
after fourth stage
Final composite strip
Same system set-up as above,
but different leach solution
used (2092, jarosited,
Cu SX prior treatment)
Starting Solution. pH • 3.0
Raffinate after second stage
contact, one hour
Raffinate after fourth stage
contact, one hour
Raffinate after second stage
contact. 3 hours
Final composite raffinate
Final Strip. 3 hours
16" Liter Test (40 v/o OEHPA)
Starting Solution. pH • 2.0
91. (Continued)
Concentration (gpl)
Zn
LJi
1JZ
SLJi
OJI
21.85
LSI
QJ2
0.04
0.24.
0.06
42.83
6.20
Fe
0.29
0.06
0.02
• o.oi
-------
TABLE 8.91. CONTINUED
.160 liter test conditions; 40 v/o DEHPA, 60 v/o KEW1AC 470B
4-stage extraction (0/A • 1)
pll adjusted after first two contacts back to pH
3-stagc strip (0/A - 1), 200 gpl H.SO.
Flowrate: 250 cc/mln. all phases
Temperature: 30-40°C
Ol
-------
8.8. SOLVENT REAGENT DEGRADATION TESTWOKK
3.8.1. Copper Solvent Extraction: LIX-622
8.8.1.1. Continuous Testwork
A series of tests were perfonned to investigate potential reagent
degradations. A large number of load/strip cycles Mere performed in the Bell
Engineering testraclc over a period of 11 days. The extraction results and a
summary of the degradation results were reported previously in Sections 8.6.3
and 6.3.4.
A small amount of organic (3.88 1 of 15 v/o LIX-622) was exposed to a
large amount of aqueous solution (>340 I) in a series of three extraction cells
and two strip cells. The smaller Bell system was chosen over the larger
Reister system so that less leach solution was required. The Reister system (5
cells) held 18.5 liters of organic (at an 0/A = 1 loading). The amount of
aqueous leach solution that could be contacted with the orga»1c in one
eight-hour day was 192 liters. Therefore, the largest possible aqueous/organic
contact ratio achievable per day was 10.4. The same aqueous/organic ratio In
the Bell system required 40 liters of aqueous solution. Therefore, the 340
liters of (aqueous solution) of contact achieved in the Bell system would have
required 3540 liters of aqueous solution in the Reister system. A decision
was. therefore, made to conduct the long-term multiple load/strip testwork in
the Bell system.
The California sludge contained primarily chromium and nickel. The sludge
was leached, then doped with copper, iron, and zinc.
The contact system was described previously. Section 5.2.1. It consisted
of three stages of extraction (loading) and two stages of strip. The
organic/aqueous ratio was one in all cells. Solution pH was adjusted before
entering the first cell but was unadjusted thereafter. Phase mixing and
settling times were both (in all cells) approximately seven minutes at a
flowrate of 50 cc/min.
347
-------
The organic Mas Ib v/o LIX-622, KS v/o KEHMAC 4701). The system cells and
pump reservoir were loaded with 3.88 liters of organic, 1.8 liters of leach
solution (cells 1.2.3) and 1.2 liters (cells 4.5} of strip acid (200 gpl
H.S04). The strip cells were fed from a four liter reservoir. The acid
contact was maintained at 200 gpl by systematic replacement of acid as.
required.
The system was started up, flows and interfaces established and sampling
of each rar'findte was conducted periodically. The system was run under
specified conditions of copper content, solution pH and temperature. Crud-
formation (if any) and pnase disengagement was observed and noted. The test
period continued until a desired volume of equeous was run through the system;
then shut down overnight. The next run was initiated the next day using a new
feed solution. The leach solution In the system from the prior day's contact
was unchanged. Interfaces were already established and it was found that
system restart involved no more than calibrating flow rates and turning on the
agitator and pump motors.
The extraction results for the eleven day test period were presented in
Table 8.84. Extraction of copper was excellent and selective. Crud formation
was unimportant and the small amount that did form was most likely fine
particulate material from the filter unit operation. It was easily removed by
aspiration.
8.8.1.2. Degradation Results
Degradation was followed by: the ability of the system to maintain low
copper content in the final raffinate; and by a special test procedure
performed on samples of organic collected at the end of each day's test.
Neither of these tests showed degradation to be occurring. The results for the
system final raffinate are presented in Tables 6.19, 6,?0, 8.81 and 8.84.
The degradation test performed on the periodic organic samples consisted
of: stripping the organic twice with unused 200 gpl H2$04 acid; contacting the
stripped organic with a stock mixed metal leach solution containing 3.11 gpl Cu
(0/A = 1); then recontacting the leach solution with another sample of stripped
348
-------
system organic (0/A » 1); and measuring the effect of the two contacts on the
copper (ind other metal) content by ICP analysis. The results are presented in
Tables 6.19 and €.20. These tables are reproduced here as a convenience to the
reader (8.92. 3.93).
Degradation of extractant for the conditions tested does not appear to be
Important. Approximately three lite-s (3.88) of organic was contacted with
over 340 liters of aqueous leach solution over a period of 113 hours. This
contact Involved approximately 226 load/strip cycles; 678 mixer contacts of
loading and 452 mixer contacts of stripping. An aqueous to organic contact
ratio of 88 was achieved for the test period; an aqueous to LIX-622 reagent
contact ratio of 587 was achieved for the test period.
Further, long term continuous testwork may be desirable but significant
deterioration by the solution conditions of high metal content, high ionic
strength, the presence of phosphorus, low pH. and mild temperature (40-50°C)
does not appear to occur. It would be desirable to conduct a detailed analysis
on the organic phase to determine if degradation of the LIX-622 oxime could b<*
followed directly. This laboratory was not capable of performing such analyses
and, therefore, an Indirect approach was chosen.
8.8.2. Iron. Zinc Solvent Extraction; D,EHPA
8.8.2.1. Continuous Testwork
A series of tests were performed to investigate potential reagent
degradation. A large number of load/strip cycles were performed in the Bell
Engineering testrack over a period of eight days. The extraction results were
reported previously in Section 8.4.2.
A small amount of organic (7.6 liters of 02EHPA) was exposed to a large
volume of aqueous solution (150 1) in a series of four extraction cells and six
strip cells; four suIfuric, three hydrochloric (Figure 8.20). The Bell system
was chosen over the larger Reister system so that less leach solution was
required. The Renter system (10 cells) held 37 liters of organic (at an 0/A =
1 loading). The amount of aqueous leach solution that could be contacted witn
349
-------
TABLE 8.92. LIX 622 LONf, TERM EXPOSURE DEGRADATION TEST SWWRY
Sample No. Organic Exposure
To Aqueous Phase
Contacts
Copper Concentration In Aqueous Phase (gpl)
System Organic
New Organic
Starting Aqueous Solution, 3.112 gpl Cu
First Day
347(1
3479
3400
3481
S 3495
3496
3497
3498
3514
3515
3516
3517
3536
3537
3540
3541
46. S liters
H *
None
n
Second Day
86.5 liters
n H
None
N
Third Day
125.5 liters
ii
None,
N
Fourth Day
161.5 liters
n ••
None
N
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
0.061
0.008
0.103
0.027
0.114
0.016
0.109
0.019
0.121
0.006
0.031
0.000
0.03S
0.016
0.028
0.001
-------
bJ
01
COKMNUED
Sample
3SSO
3551
3615
3616
3617
3616
363S
3636
3637
3638
3647
36<8
3649
3650
3665
3666
3667
3668
No. Organic exposure Contacts Copper Concentration In Aqueous Phase (gpl
To Aqueous Phase,
Fifth Day
167.0 liters
None
Sixth 0«y
206.5 liters
None
H
Seventh Day
241.0 liters
M N
hone
»*
Eighth Pay
275.5 liters
i i
(lore
Ninth Pay
287.5 liters
II M
None
ii
First
First
First
Second
First
Second
First
Second
First
Second
First
Sue oud
First
Second
First
Second
First
Second
System Organic
0.001
0.036
0.008 '
0.112
0.022
0.258
0.023
0.120
0.000
New Organic
0.040
0.006
0.007
0.024
0.007
0.051
0.004
0.015
0.020
-------
CONUNUED
Notes: . Conditions for each days exposure given (n Table 8.8?.
. Degradation test conditions: 50ce system organic stripped twice (0/A -1)
with unused 200 gpl I^SOd.; stripped organic contacted with copper stock
solution. pH • 1.36 for first four tests, pl< » 2.0 for last five tests; a
second system organic sample contacted same stock solution, I.e., stock
solution was contacted twice with two used organic samples, stock pll not
adjusted between contacts.
Unused organic sane as system organic, IS X LIX 622. contacted with a 30
gpl Cu, 200 gpl H2S04 solution, then contacted with slock solution as
... described above for system organic.
U1
-------
Ul
ill
ut
Sample No.
TABLE 8.93. LIX 622
LOADING
Organic Exposure
To Aqueous Phase
LONG TERM
EXPOSURE DEGRADATION TEST SUMMARY:
Contacts Loading,
System Organic
Stock Aqueous Solution, 3.112 gpl Cu, 3.958 gpl Fe. 0
2.014 gpl Cr. 6.061 op) Ni. 0.287 gpl Al, 0.319 gpl
3478
3479
3480
3481
3495
3496
3497
3496
3514
3515
3516
3517
3536
3537
3540
3541
First Day
46.5 liters Aqueous
• H
None
N
Second Day
86.5 liters
M H
None
u
Third Day
125.5 liters
•
None
M
Fourth Day
161.5 liters
H
None
II
First
Second
First
Second
First
Second
Mrst
Second
First
Second
First
fecond
First
Second
First
Second
0.203
0.004
0.200
0.005
0.200
0.006
0.200
0.006
gpl/S LIX 622
New Organic
.122 gpl Zn.
Ca
0.199
0.008
0.205
0.014
0.705
0.001
0.206
0.002
-------
CON1IIIUED
Sample No.
3550
3551
3615
3616
3617
3618
3635
3636
3637
3638
3647
3648
3649
3650
3665
3666
3667
3663
Organ 1: Exposure
To Aqueous Pnase
Fifth Day
187.0 liters
None
Sixth nav
206.5
•
Hone
•
Seventh Day
241.0 liters
•
None
M
Eighth Day
275.5 liters .
n
None
M
Ninth Day
207.5 Uteri
•
Hone
M
Contacts
First
First
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
Firs:
Second
loading,
System Organic
0.207*
0.205
0.002
0.200
0.006
0.190
0.016
0.199
» 0.013
gpW 1 IX 622
New Organic
0.205
0.207
0.206
0.001
0.204
0.003
0.206
Note: . Conditions for each days exposure presented In Table 8.83.
-------
the organic in one eight-hour day was 192 liters. Therefore, Che largest
possible aqueous/organic ccitact ratio achievable per day was 5.2. The same
aqueous/organic ratio in the Bell system required 40 liters of aqueous
solution. Therefore, the 150 liters (of aqueous solution) of contact achieved
in the Bell system would have required 780 liters of aqueous solution in the
Reister system. A decision was. therefore, made to conduct the long-term
multiple load/strip testwork in the Sell system.
The contact system is shown schematically in Figure 8.20. It consisted of
one stage of low pH contact for high iron removal from the leach solution; and
three stages of iron and zinc extraction; one stage of zinc stripping (by
sulfuric acid) from the Iron loaded organic followed by three stages of Iron
stripping by hydrochloric add; and two additional stages of zinc stripping by
sulfuric acid. The organic/aqueous ratio was one in all cells. Solution pH
was adjusted to approximately one before entering the first cell, then
readjusted to approximately two before entering the second cell. The solution
pH was unadjusted thereafter. Phase mixing and settling times were both (in
all cells) approximately seven minutes at a flowrate of 50 cc/min.
The organic was 40 v/o D-EHPA, 60 v/o KERMAC 510. The system cells and
pump reservoir was loaded with 7.6 liters of organic; 1.8 liters of leach
solution (cells 1.2.3.4); 2.4 liters of sulfuric strip acid (200 gpl H2$04,
cells 5,6.7); and 2.4 liters of hydrochloric strip acid (6 N HC1, cells
8,9.10). The a:id contents of the strip cells wera maintained at their desired
strength by systematic replacement of acid as required.
The system was started up, flows and interfaces established and sampling
of each raffinate was conducted periodically. The system was. run under
specified conditions of zinc content, solution pH and temperature. Crud
foi-mation and pnase disengagement was observed and noted. The test period
continued until a desired volume of aqueous was run throug the system; then
shut down overnight. The next run was initiated the next day using a new feed
solution. The leach solution in the system from the prior day's contact was
unchanged. Interfaces were already established and it was found that system
restart involved simply calibrating pump flows and turning on the agitator and
pump motors.
-------
Raltliiata
91
S:<
•^M
V
SET
Sta
MI:
<
t
)
FLCR
Strip
g, 1
KER
54~
k
J
>
HCI
-M
Tgure 8.20. Schematic of Iron, tine long term test system flow pattern
-------
The extraction results for the eight-day test period were presented in
Table 8.53. Extraction of iron and zinc were excellent. Crud formation In the
first cell was a problem (see Section 8.4.3) until the system diluent was
switched to KERMAC 510 kerosene.
8.8.2.2. Degradation Results
Degradation was followed by: the ability of the system to maintain low
Iron and zinc concentrations In the final raffln&te; and by a special test
procedure performed on samples of organic collected at the end of each"day's
test. Neither of these tests showed degradation to be occurring. The results
for the system final raffinates are presented in Table 8.53.
The degradation test performed on the periodic organic samples consisted
of: stripping the organic twice with unused 200 gpl H_S9. add; contacting the
stripped organic with a stock mixed metal solution containing more iron and
zinc than a 40 v/o 02EHPA organic could extract containing 11.64 gpl Fe, 11.19
gpl Zn (0/A • 1); then recontactlng the leach solution with another sample of
stripped system organic (0/A - 1); and measuring the effect of the two contacts
on the zinc and Iron (and other metals) content by ICP analysis. The results
were presented previously In Table 6.24 and are reproduced here as a
convenience to the reader (8.94).
Degradation of extractant for the conditions tested does not appear to be
Important. Approximately 7.6 liters of 40 volume percent DgEHPA-60 in KERMAC
510 kerosene was contacted with over 150 liters of aqueous leach solution over
a period of 67 hours. This contact Involved approximately 58 load/strip
cycles; 232 mixer contacts of loading and 586 mixer contacts of stripping. An
aqueous to organic contact ratio of 20 was achieved for the test period. An
aqueous to O.EHPA reagent contact ratio of 50 was achieved for the test period.
Degradation r•-.-'. results are presented In Table 8.95. No noticeable
decrease in the ib'.'.izy of the reagent to effectively extract iron and zinc is
shown over the test period considered. It would be desirable to conduct a
detailed analysis on the organic phase to determine if degradation of the
357
-------
TABLE 8.94. DEIIPA LONG TERM EXPOSURE DEGRADATION TEST
Sample Ko.
4025
3841
3842
3843
3844
3874
3875
3876
3877
3909
3910
3911
3912
3947
3948
Organic Exposure
To Aqueous Phase
Stock Aqueous Solution,
pH - 2.0. 11.639 gpl Fe,
11.192 gpl Zn
First Day
19 liters aqueous
None
k
Second Day
3C liters
None
Third Day
57 liters
H N
None
Fourth Day
76 liters
II II
Contacts
i
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
First
Second
Loading, gpl
System Organic
0.257
0.068
0.261
0.054
0.242
0.066
' 0.271
0.056
(Zn«re)/t OCHPA
New Organic
0.327
0.032
0.270
0.056
0.286
0.051
-------
VO
CONTINUED
Sample No.
3947
3948
3982
3983
3984
3985
4026
4027
4028
4029
Organic Exposure
To Aqueous Phase
None
N
Fifth Day
96 liters
• M
None
M
Sixth Day
115 liters
• N
None
M
Contacts
First
Second
First
Second
First
Second
First
Second
First
Second
Loading, gpl
System Organic
0.248
0.054
0.286
0.029
-------
TABLE 8.95 LONG-TERH DEGRADATION-EXTRACTION STUDY: ZINC AND IRON BY 401 DEHPA
01
o
Sample Ho. Conditions
3798
3835
3871
3908
Aqueous Exposure
to Organic
t Hrs.
First Day
19 14
Second Day
38 23
Third Day
57 29.5
Fourth Day
76 37.5
"Mixer Contact Time, roin.
Feed Strip
336 252
552 414
708 531
900 675
3944
Fifth Day
96 44.5 1068 800
I'd
252
414
531
675
Concentration in Final Cell 4 Composite
Raffinate. gpl
800
Fe
0.342
(high because of
incomplete iron
oxidation)
0.070
O.C30
0.319
(high because of
incomplete iron
oxHitlon)
0.027
Zn
0.005
0.094
0.106
0.046
0.031
-------
CONTINUED
Concentration In Final Cell 4 Composite
Mo. Conditions Reffinate. gpl
Aqueous Exposure
to Organic *Hlxer Contact Time, nin.
• Mrs. Feed Strip fe Zn
HC1
u» Sixth Pay
~* 3975 115 51.& 1230 927 927 0.238 0.046
(high because of
incomplete oxidation)
Seventh Day
4022 133 59 1416 1062 1062 0.051 0.066
Eighth Day
4C»4 151 67 1608 2010 2010 0.028 O.IK 3
NOTES: 'Conditions for each day's exposure presented In Table 6.83.
'Mixer contact time is tine organic was exposed to feed solution or to strip solution.
-------
DgEHPA reagent (or if decomposition products could be identified) could be
followed directly. This laboratory was not capaole of performing such analyses
and, therefore, an indirect app-oach was chosen.
8.9. CHROMIUM OXIDATION
Selective removal of Cr ions from a mixed metal solution containing
Cu*2, Fe*2'*3, Zn*2, Mi*2, and Al*3, does not appear possible without
-2 -2
conversion to an oxidized, CrO. or Cr.O. , form. Conversion to chrcmate or
dichromate requires a strongly oxidizing environment which means that the
oxidation must be accomplished after solvent extraction processes because a
strongly oxidizing solution is expected to degrade the organic extraction
reagents. The proper place for the chromium oxidation unit operation,
therefore, appears to be after Fe, Cu, and Zn have been removeJ.
There does not appear to be a successful way of selectively separating
Cr from mixed metal solutions by SX. Reagents that have been successful in
extracting Cr from an aquecus phase suffer from slow strip kinetics when
using H.SO. acid or require that HC1 be the stripping agent. Certainly
+3
consideration of Cr removal by SX techniques Is in the laboratory stage of
investigation rather than in industrial practice. Known studies on Cr* SX a: e
listed in Reference 40.
Two approaches to chromium oxidation are proposed, i.e., 1) solution
oxidation or 2) sludge oxidation by roasting. Each approach has several
alternate means available to accomplish the oxidation.
8.9.1. Solution Oxidation of Chromium
Emphasis was placed in this study on solution oxidation techniques for
chromium. The experimental technique used to determine the degree of oxidation
wai: expose a known volume of solution to a set of oxidizing conditions;
separate solids if present; treat a known volume of the solution to ion
exchange using the anion exchanger IRA 900 (this quantitat' »ly removes all
anIon species formed during the oxidation, i.e., Cr04°, Cr^O-j", or
releach any solid phase to complete the mass balance.
362
-------
Studies have been conducted using H^O.. PbO-, Na-O-, Cl_, and aqueous
cnloride ion containing species such as HC10. The emphasis was placed on Cl-
as the oxidizing agent became of its ability to easily raise the solution Eh to
+3 -2 2
above the Cr /CrO. (or Cr_0, ) equilibrium half cell potential.
8.9.1.1. Chlorine Oxidation of Chromium
8.9.1.1.1. Phase I Studies
Small scale testwork was conducted on: chromium solutions made in the
laboratory; chromium bearing leach solutions procreated in a variety of ways;
a-id chromium bearing leach solutions produced in the large test assembly. The
experimental results are presented in Tables 8.96-8.98.
All testwork was performed on leach solutions that had been pretreated for
removal of most of the iron, copper, and zinc, i.e., a leach solution was
oxidized that contained primarily chromium and nickel. Preliminary testworlc
showed that oxidation rates were significantly more rapid in slurry bearing
solutions (4 1000 mv, and pH was maintained at a specified value (Table
8.97); or chlorine was sparged into a pH adjusted (pH = 5) solution, without
maintaining the pH, until all solids were dissolved, then the cycle repeated
(Table 8.98).
The results presented in Table 8.96 show a comparison of the chromium
oxidation achieved in three types of leach solutions. The chromium oxidation
In a specific time period is less in unpretreated leach solutions (containing
Cu. Fe, Zn, Cr, Ni, Cd, and Al) than in jarosite treated (Fe removed) and
jarosite-CuSX-ZnSX (Fe, Cu. Zn removed) treated solution?.
•
3C3
-------
TABLE 9.96. CHLORINE OXIDATION Or CHROMIUM AS A FUNCTION OF PH AND TYPE
STARTING SOLUTION
Sample No. Condition Chromium Oxidized (")
1964 Leach Solution (2.44gp1 Cr*3)
2009 pH ' 2.2 58.3
1979 pH • 4.0 59.7
pH - 5.0 63.8
1988 Jarosite Treated Leach Solution (S.SOgpi Cr*3)
2011 pH • 2.1 60.0
2027 pH • 4.0 85.8
2033 pH - 5.0 84.9
2038 Jarosite - Cu SX - Zn SX Treated Leach Solution (4.46gpl Cr*3)
2051 pH - 1.8 47.5
2040 pH - 4.0 85.9
2046 pH - 5.0 83.3
Notes: . All oxidation tests performed on 103 cc of pH adjusted leach
solution; pH maintained; 0.2 liters Cl./ min.; contact time
one hour.
. Analytical procedure: solution contact performed; sample
filtered; solution analyzed; solution contacted with IRA 900
for chromate removal; solution from IX analyzed; solids
leached: all solution volumes recorded; oxidation calculated
from mass balance.
. Starting solutions prepared by standard leach of barrel 12
sludge. The starting solution concentrations (gpl) were:
_Cr_JH__Al_ _Fe__Cu__Zn_ Cd
1964 2.44 4.55 1.52 8.32 0.73 2.91 0.23
1988 5.80 5.60 1.66 0.73 0.84 2.99 0.37
2038 4.46 3.91 0.74 0.03 0.02 0.29 0.24
364
-------
TABLE 9.97.CHLORINE OXIDATION OF CHROMIUM: Eh MAINTAINED
Sample No. Condition Chromium Oxidized (I)
2190 Starting Solution (3.00gpl Cr*3)
pH • l.S. Eh - 374mv
pH • 1.4
2200 Fifteen orinute exposure.ll60mv 8.6
2206 One hour exposure. 1140mv 7.3
pH - S.O
2234 One-half hour exposure. 1138ow 81.0
2237 one hour exposure. 1136nv 84.6
2181 Starting Solution (0.37qp1 Cr*3)
pH • 1.5. Eh. • 191mv
pH • 1.5
2197 one hour exposure, 1135iw 10.3
pH • 4.0
2212 One-half hour exposure. 1088mv 30.5
2215 One hour exposure. 1in2mv 45.8
pH • 5.0
2223 One-half exposure. 1088tr,v 54.8
2226 One hour exposure, 1099mv 72.7
Notes: . Starting Solution Composition (gpl):
Cr N1 A1 Fe Cu Zn Cd
2190 3.00 2.39 0.46 0.01 0.08 0.12 0.23
2181 0.37 2.54 0.49 0.01 0.08 0.13 0.24
. SOOcc solution from sequential test series four. Solution treated
previously for Iron, copper and zinc removal. Series 2'90 was
doped with chromium.
. Chlorine addition rate 0.2 1/rrin but only supplied -erlodlcally
to keep the solution Eh >1000im». PH maintained.
365
-------
TABLE 9.98. CHLORINE OXIDATION OF CHROMIUM: CYCLE TEST
Saople No. Condition Qirox'um Oxidized (1)
2417 Starting Solution (2.47gpl)
pH • 1.43. Eh » 373mv
2418 pH adjusted to 5.0. chlorine 40.0
purged at 0.2 I/mm until
all solids dissolved.
Eh>- lOOOmv, 25 nrin.. SCO cc
2421 pH readjusted to 5. above 87.7
procedure repeated, 10 min.
Eh» llOOfflv
2424 pH readjusted to 5. above 88.7
procedure repeated. 5 min.
Eh* llOOmv
Notes: . Star-tint) solution conposition (gpl): 2.93 II:, 2.<7 Or. 0.06 Fe.
0.11 Zn. 0.09 Cu. 0.28 Cd. 0.04 Al
. 500 cc solution from sequential test series four. Solution
treated previously for iron, copper and zinc removal.
Solution doped with Cr to achieve reported conrentration.
. pH adjusted to S then chlorine added at 0.2 1/min uni.l pH
was lowered and the solids all redissolved. Procedure
repeated three times.
366
-------
The results presented in Table 8.47 show that rpasonably effective
chromium oxidation occurs at d pH of b if the Eh of the solution is maintained
> 1000 mv. Poor chromium oxidation is achieved at lower pH values. The
results presented in Taole 8.98 show that the rate of chromium oxidation -nay be
greater if a cyclic precipitation-oxidation sequence is followed.
Large scale testwork resulted in effective slurry oxidation of chroipiunn,
I.e., >80% conversion. All unoxldized Cr*+* remains as chromium hydroxide and
can be conveniently recycled to the original leach unit operation. The results
of the large scale testwork are presented in Table 8.99. The 30-liter test
resulted in producing a solution (1.65 gpl Cr) that was almost completely
oxidized (95*). The 75-liter test resulted in producing a solution (83.6% of
the chromium oxidized) that contained 2.28 gpl Cr (95.2% Cr*6, 4.8% Cr*3).
Most of the unoxldized chromium remaining In the system after oxidation
was present as solid Cr(OH).. This solid can be recycled to the leach system
and. therefore, does not represent a loss of chromium.
The large scale testwork showed a greater time required for effective
oxidation than the small scale tests. The main reason for this is that the
sparging system in the large scale testwork did not produce as efficient
gas-so'ution-solid contact interfaces. It was thought that this difficulty
could be overcome by modifying the sparging reactor design. This did not,
however, prove to be true.
8.9.1.1.2. Phase II Study
Additional large scale testwork was performed on chlorine oxidation of
chromium during the Phase II study. The use of a chlorinator was Investigated.
Such devices find widespread use in many industries.
A chlorinator in its simplest design resembles an aspirator commonly used
In laboratories; a schematic drawing is presented in Figure 8.21. Liquid is
pumped through a venturi orifice. The solution flow creates a low pressure at
a side port. Chlorine Is sucked into the aspirating chamber through the side
367
-------
TA3LE 8.99. CHROMIUM OXIDATION IN
Sample No. Condition
£
00
2181a
2361
2340
2347
2564
2574
JO Liter Test
Starting Solution before
pll adjust (pll * 1.3, Eh •
380 "w »
Starting solution adjusted
to pH = 5.0
fine hr sample, exposed
periodically to C1-;
Eh maintained at
1000 tnv and pll 4.
A portion of above solution
13. C lit.) re-exposed to
flowing C!2, 1.5 hr
7«i Liter Test
Starting Solution before
pH adjust
Two hr exposure
Cu
0.08
< D.L.
0.04
LARGE SCALE SEQUENTIAL TESTHORK
Concentration (gpl)
Fe
0.01
0.001
0.01
. Zn Cr
0.13 '..69
< D.L. 0.27
0.08 1.15
Hi
2.54
1.75
2.14
Cd
0.24
0.23
0.2S
Al
0.4)
95rox"idized)
0.04
0.03
0.22
0.08
0.07 2.67
0.05 1 . 70
1.75
1.45
0.11
0.09
U.39
0.19
(SZT'^oxid.)
,.•589
Five hr exposure
0.03
D.L.
0.06 ?.?8
1.68
0.11
0.08
(83.6T~oxid.)
.tote: . Oc Lit led experimental results presented In Table 0.125, 8.127.
-------
Solution
Discharge
\\X\\\\\\ \ VN.
\\\\\\\\\\yvx
o
Injector
Inlet
Chlorine Gas
Figure 8.2). Schematic drawing of chlorlnator.
-------
port. Turbulence is created in the liqu:d and good gas (chlorine)-liquid
contact is achieved.
Chromium is more effectively oxidized in a slurry where the chromium is
present as chromium hydroxide (pH = 4-5):
7 * 6 N*0" - > 2 Cr(OH), + 3 Na-SO,
3 3(solid) Z 4
2 Cr(OH)3 - > 2 HCr02 + 2 H20
2 HCr02 + 4 H2C + 3 C12 - > 2 HgCrO^ + 6 HC1
Hydrochloric acid is generated and must be neutralized with caustic so that the
solution alkalinity remains at a pH of 4-5.
Some nickelous hydroxide may be formed as a solid under the chlorinating
conditions but the quantity is small and the residual chromium hydroxide and
the nickelous hydroxide can be recycled to the original leach unit operation.
A photograph of the oxidation system is presented In Section 8.14. A
schematic representation of the system is presented in Figure 8.22. The system
consisted of two 100 liter tanks. Tank A and B were connected so that solution
slurry could flow easily from tank B to tank A. Tank A was the feed tank; the
slurry solution in each tank was agitated. The slurry in Tank A was pumped to
the chlorinator. Discharge from the ch'orinator flowed into tank 8. Solution
pH was maintained in tank B by a pH controller. Shut-off valves placed on the
Inlet and discharge lines prevented loss of liquid from the piping.
Each oxidation test was begun by placing 120 liters of chromium and nickel
containing leach solution in tanks A and B. Agitators were turned on. The pH
controller was set with a low limit pH value so that when the pH value was
reached or sensed, the controller would activate the small feed pump to add
caustic to the tank.
370
-------
u>
Agile-Tor
Tent. I
Figure 6.22. Schematic drawing of the chlortnator system.
-------
Severe! shakedown tests were performed to observe operations and to
Identify potential problems. Flow problems, plug-ups. and contact problems
were noted and corrected.
The following sampling technique was used:
'Samples of slurry were taken from tank B periodically as a function
of time.
'The sample was vigorously agitated and an aliquot was taken.
'The aliquot was acidified with a known amount of sulfuric acid.
'Total Cr and Ni were analyzed in the acidified solution (solids had
redissolved).
"The original sample was allowed to stand and solids to settle.
'All aliquot of the solution was taken. This sample represents the
filtrate r-f the chlorine oxidation operation in an actual
solid-liquid separation that would be carried out commercially.
Chromium and nickel were determined in the liquor.
*A second solution aliquot was taken from tne settled sample. The
liquid was exposed to Rohm and Haas ion exchange resin 1R-900 (an
anionic exchange resin).
Chromium anions are extracted (exchanged for S0.~ ions); chromium cations
(Cr***) are not extracted. Therefore, by measuring the effluent solution for
chromium the concentration of oxidized chromium can be determined by difference
between the total chromium going into the exchange resin and the chromium
carryout.
The results of the large scale chlorination testwork are presented in
Tables 8.100-8.102. The results show only up to about seventy percent chromium
oxidation in a nine hour period exposure. The previous testwork using a
sparging system resulteo in oxidation conversions of over eighty percent in
four hours. At this point it appears thct electrochemical oxidation is the
more appropriate approach to follow.
372
-------
TABLE 8.100.
CHLORINE OXIDATION OF CHROMIUM: LARGE SCALE TEST
Exposure Tine Slurry (gpl) Decant (gpl)
Oxidation (1)
Cr10"1 Ni IOLal CrrotalJ>*J
Starting
1 Hr
2
3
4
2.02
2.02
2.02
2.02
2.34
3.38
3.16
3.02
3.24
2.98
1.06
1.25
1.48
1.45
1.45
0.38
0.42
0.30
0.23
O.il
33.7
41.1
58.4
60.4
57.3
Notes; .pH maintained ati 4.5.
.Starting slurry was from a previous shakedown test.
.Chlorine feed rate O.i7 lD/hr.
373
-------
TABLE 3.IO:. CHLORINE OXIDATION OF CHROMIUM- TLiT TWO
Exposure Time Slurry (gpl )
1 Hr
2
3
4
S
6
7
8
9
Re-erpos<"J
1
2
3
4
5
10
, Total
Cr
2.72
2.52
2.52
2.72
2.58
2.86
2.66
2.80
2.66
above solution
3.94
4.15
4.21
4.27
3.94
4.17
TOtdl f lO^c
Ni Cr
5.84
5.32
5.70
5.70
5.08
5.46
4.83
4.83
4.83
?9
.05
.12
.40
.54
.en
.81
.89
.81
. pH raised from
4.83 3.09
4.37 3.06
4.70 2.97
5.06 2.92
4.70 2.92
5.06 2.86
Decant (gpt;
H r -3
Cr
0.75
3.57
3.44
0.20
0.17
0.05
b.03
0.04
0.03
4.5 to 5 S.
0.03
0.03
0.02
0.03
0.03
0.02
Oxidation ('-)
19.3
19.0
27.0
40.8
53.1
57.0
67.0
66.1
67.0
TEST T"REE
77.7
73.0
69.8
67.7
73.3
68.0
Notes;
Chlorine feed rate 0.42 Ib/hr.
See text for details
374
-------
TABLE 8.102. CHLORINE OXIDATION OF CHROMIUM: fEST FOUR.
Exposure Tine
Starting
Solution
1 Mr
2
3
4
5
Slu'ry (gpl)
CrTotal
4.24
4.52
4.57
4.46
4.46
4.52
Ntal
0
0./6
1 24
1.47
2.09
2.45
Decant (qpl
Cr'3
0
0.02
0.03
0.03
0.05
O.US
I Oxidation (X)
16.4
26.5
32.3
4S.7
53.1
Notes: . pH maintained at • 5.5
. Chlorine feed rule at maximum race for system. 0.42 Ib/hr.
375
-------
8.9.1.2. Electrochemical Oxidation
8.9.1.2.1. Phase I Study
Chromium can be electrocnemically oxidized in a compartment!zed
electrolytic cell. The technology has been developed by the U.S.B.M.* ' a*
Rolla, MO, and is commercially available, e.g.. Scientific Control
Laboratories. A schematic diagram is presented in Figure 8.23. The system
consists of a series of cells made up of an anode chamber where Cr is
oxidized to Cr.O^* and a cathode chamber where metal ions are deposited. The
chambers are separated by ion selective membranes such as OuPont Nafion 423.
These membranes are cation selective, i.e., they allow only cations to pass
Into the cathode chamber. The separation membrane is necessary in order to
prevent oxidized chromium frow being reduced at the cathode.
Pilot studies by U.S.B.M. have shown successful oxidation of Cr at
energy consumptions of 9 kw hour per kg Na-Cr-O, produced.
Preliminary experiments have been performed in the Montana Tech laboratory
to study the potential for application of thi«. oxidation technique to the
t/^esent type of solutions. A schematic diagram of the test cell is presented
in F'gure S.i'j. Th* preliminary experimental results ar-e presented
in Table 8.103.
8.9.1.2.2. Phase II Study
Electrochemical oxidation of chromium wstun are included in Section 8.14.
The electrochemical cell consisted of two box chambers capable of treating
about 14 liters of anolyte and 26 liters of cathulyte in a batch or continuous
mode of operation.
376
-------
""
o
•*»«««
SF€NT
XXUT1ON
1
/>
N
HtGCNSUTEO
toumoH
s>
aCCTHOLVTA
Figure 8.23. Electrochemical cell for oxidation of chromium. (Supplied
by Scientific Control Laboratories, Inc.)
377
-------
TABLE 8.103. ELECTROCHEMICAL OXIDATION OF CHROMIUM: PRELIMINARY
TESTUORK
Sample No. Condition Chromium Oxidation Extent (S)
Batch Test
'266'. Starting Solution (2.90 gpl Cr*3)
Exposure of 1.1 liters
of solution to 3 volts at,
8 amps. C.O. • 20 amps/ft .
2665 Three hour exposure 69.9
2667 Four hour exposure 84.3
2670 Five hour exposure 89.3
Continuous Test
2664 Smarting Solution (2.90 gnl Cr*3)
Exposure of 1.1 liters
of solution flowing at
3-5 ce/fflln to 3.5 volts at
12 amps. C.0.« 20 amps/ft*.
2673 One a.id one half hour exposure 69.5
2683 Three hour exposure 75.4
2689 Six hour exposure " 81.6
Notes: . Cell description presented In Figure 5.8.
. Starting solution was zinc ruffinate from the large scale
test solution, sequential test series five. Iron, Cu, and were
removed prior to use in this study.
. Batch Test Conditions: The cell was loaded with: anolyte -
1.1 liters of leacn solution, catholyte -
2.3 liters of ISO apl H?S04.
The solution was exposed for a period of tiers, then sampled. The
solution was subjected to an anion exchange re
-------
entrapped solution and this solution was anatyjed for chromium.
The oxidized chromium was then calculated from the mass
distribution data.
Continuous Test; The eel I was loaded wi th
anolyte - 1.1 liters nf composited batch test solution.
catholyte - 2.3 liters of 180 gpl H^O^.
The solution was fed for 1.5 hours at 5 cc/rtn then for
3 hours at 3 - 4 cc/rrin. Solution was sampled as a function
Of time and analyzed as described above. The chromium
in tie composited final solution was 75.3 1 oxidized; the
pH MBS 0.33. Eh was 866 mv.
379
-------
Initial tests were conducted using an anoae:diapnragm:cathode surface area
ratio of 1:1:1. Applied voltage was 3.5v-4.5v. current density was 8-20
amp/ft'. Static batch test results are reported in Table 8.104. Approximately
85% oxidation was achieved within 21 hours. Further exposure tiad no apparent
effect.
Three continuous tests were performed. The solution flowrates were 10
cc/min. for additive leach solution and discharge anolyte. The catholyte was
withdrawn and recirculated back into the catholyte chamber. A perforated lead
wool anode was used to ensure a high contact surface area. Lead sheet cathodes
were used. The anode surface area available to the solution is unknown, the
o
physical outer holding plate area was 1.8 ft . The cathode surface area was
1.32 ft2. . -
The results of the first continuous test series are presented in Table
8.105. The anolyte chamber was filled with fully oxidized chromium; unoxidized
leach solution was fed into the chamber at a rate of 10 cc/min. The effluent
stream showed 81" of the chromium oxidized.
A second test series was run for 48 hours using the anolyte from Series
One, Table 8.106. The conversion achieved in this test was 87.2%. A third
series of tests were conducted using u new catholyte. Table 8.107. Steady
state conditions appear to be established after about 48 hours at a conversion
rate of about 901 of the chromium.
The Nafion membrane allows cations to pass through but not anions. If the
assumption is made that no solution leaked from the anode chamber to the
cathode chamber then migration of nickel can be determined. A determined
effort was made to prevent leakage, therefor2, the assumption is probably
reasonable. If Cr had been analyzed in the catholyte then leakage could have
been detected. This analysis was, however, not performed. The nickel
migration results are presented in Table 8.107.
8.9.1.3. Oxidizing Properties of SO--U- Systen
An alternate (perhaps low cost) means of oxidizing chromium may be by a
relatively new technology developed by INCO* ', i.e., an S02-02 system. INCO
380
-------
TABLE 8.
.104 BATCH ELECTROCHEMICAL OXIDATION OF CHROMIUM
Sample No. Condition Chromium Concentration (gpl )
50008
5001
5002
5003
Anolyte
Cr*6 Cr*3
Starting Solu. O.U 2.91
21 Hrs 2.72 0.27
24 2.73 0.23
48 2.54 0.23
Catholyte Chromium Oxidized (X)
Crfucal
0.16
0.22 84.9
0.20 86.4
0.42 85.4
Notes: . Starting anolyte (pH > 1.0) and catholyte rrom a previous
test. Catholyte 180 gpl H2S04.
. Perforated anode. 4.5 volts. 14.5 - 19.0 amps/ft? .
. Solution air agitated.
TABLE 8.1 OS. CONTINUOUS ELECTROCHEMICAL OXIDATION OF CHROMIUM: SERIES OHL
Chromium Chromium
Sample No. Condition Concentration, gpl Oxidation. I
5004 Starting Anoly'e 2.77
SOOS Feed Anolyte. 6A lit. — 1.54
5006.7 120 Hours 1.42 0 34 0.61 78.0
NOTES: 'A « anolyte. C - c-nholyte
• Anolyte feed pH -• 1.0 and withdrawn continuously at 10 cc/rain.,
- catnolyte recirculated at 10 cc/min.. air agitation used.
•Perforated lead anode. 4.S v. 16.7 - 18 amp/ft2.
•Catholyte contained 180 gpl H2SCV
381
-------
TABLE 8.106. CONTINUOUS ELECTROCHEMICAL OXIDATION OF CHROHIUM: SERIES TWO
Chromium Chromium
Samole No. Condition Concentration, qpl Oxidation. "
5009 Starting Anolyte 1.42 0.34
5010 Feed Anolyte — 1.79
5011.12 24 Mrs. 1.22 0.35 0.23 80.4
5013.14 48 Hrs. 1.70 0.23 0.24 87.2
NOTES: -Conditions same as presented in Table 8.105.
382
-------
TABLE 8.107. CONTINUOUS ELECTROCHEMICAL OXIDATION OF CHROMIUM- SERIES THREE
Sample No
5024
5025
5023
5026,7
5028
5029
5030
5131,32
5033,34
5037,38
5039.40
. Condition
Starting Anolyte
Catholyte
Feed Anolyte
24 H-s.
Diaphragm Washed
New Feed Anolyte
New Anolyte
Catholyte Unchanged
24 Mrs.
48 Mrs.
96 Hrs.
12Q Hrs.
Chromium
Concentration, gpl
A C
Cr*6 e£3
13.95 0.15
— 2.40
11.36 0.44
2.73
25.3
19.8 0.45
13.9 0.30
9.76 0.35
8. 04 0.12
Cptotal
0.01
0.01
0.12
0.12
0.30
0.35
0.61
0.71
Chroffllum
Oxidation. I
81.7
83.5
89.0
87.1
?5.6
NOTES: -A - onolyte. C * catholyte
•Anolyte feed !pH * 1.0) continuously at 10 cc/min.. cathnlyte
reelrcuIdted but nut withdrawn.
•Perforated 'ead anode. 4.5 v. 22.5 amp/ft2.
383
-------
TABLE 8.108. NICKEL MIGRATION THROUGH NATION DIAPHRAGM DURING SERIES
THREE CHROMIUM OXIDATION TESTWORK
Simple No.
5031.32
5033.34
5035.36
5037.38
5039.49
Condition
24 Hr. Exposure
48 '
72 "
96 "
120 "
Nickel
During Period
15.9
C.6
11. 1
6.6
6.3
Migration, 7
tumulat ive
15.9
24.5
35.6
42.2
48.5
NOTES: •Conditions given in Table
•Starting nickel concentration in anolyte: 0.86 gpl
•Feed nickel concentration in anolyte: 2.P6 gpl
,Start1ng nickel concentration in catholyte: 0.03 gol.
384
-------
studies have sfc. form and serve as-
the oxidizing specie. The pH level required depends on what metal ion is to be
oxidized. Chromium oxidation has not been investigated. However, if the
system were applicable to chromium a rather low cost oxidizing system may be
possible; i.e., perhaps a considerably lower cost than the chlorine oxidizing
system or electrochemical oxidation system.
The SO.-O- approach is seen to have potential possible future application
for oxidation of chromium but the technology is not at present demonstrated for
chromium oxidation and is, therefore, not considered as a viable alternate for
the present study. The system is, however, included in the cost analysis.
Section 8.15.
8.9.2. Sludge Oxidation by_ Roasting
Oxidation by roasting may be the only feasible approach for high iron-high
chromium sludge materials because of the chromium loss during iron removal by
jarosite precipitation. The concept of oxidation roasting is that Cr is
+fi 9
converted to Cr (as Cr04 ) which reacts in the presence of sodium to form
sodium chromate. Sodium chromate is soluble at all pH values, therefore, it
can be dissolved by a water leach in preference to all other metals.
Only a few preliminary tests have been conducted during the present study.
The results are encouraging and indicates a possible research direction for
high chromium sludge material. The preliminary test results are summarized in
Tables 8.109. and 8.110.
8.10. CHROMIUM EXTRACTION
Chromium is present in mixed metal sludge leach solutions in relatively
low concentrations: usually <2-4 gpl. Therefore, a means of concentration (in
addition to selective removal) is required.
385
-------
TABLE 8.109.
Sample *
1237
1238
1239
1240
1241
1242
1243
1?44
1245
1246
ROAST-LEACH TEST ON MICH CHROMIUM
Condition
200°C Roast
1 hr.
4 hr.
4 hr.. fOOH
400°C Roast
i hr.
4 hr.
4 hr.. NaDH
600°C Roast
1 hr.
4 hr.
4 hr.. NaUl
Untreated Sludge
SLUDGE: PRELIMINARY TESTS
C.- Extraction (Z)
18.9
10.8
32.6
4.6
10.2
90.6
4.3
1.4
97.3
63.6 .
NOTE: -Starting Solid Composition (I): 1S.64±.0.46 Cr. 0.7110.05 Fe;
0.712).02 Cu. 1.1610.04 Hi,
0.061.02 Zn. Cd
-------
TABLE 8.110. ROAST LEACH ON HIGH CHROMIUM SLUDGE: TESTUORK
Condition Cr £At.raction
400°C Roast. 2 Hours
5 gn sludge. 5 gm NaOH. 39.2
2.5 gm
12S3 S gm sludge. S gm NaOH 10. 0
1264 10 gm sludge. 10 gm NaOH 72.9
600°C Roast. 2 Hours
1255 5 gm sludge. 5 gm NaOH 71.4
2.5 gm
1266 5 gm sludge. 5 jm NaOH 84.2
1267 10 gm sludge. 5 gm NaOH 37.2
800°C Roast, i Hours
1269 5 gm sludge. 5 gm NaOH 50.4
1270 10 gm sludge. 10 gm NaOH 62.4
1355 Untreated solid sludge 0.6
Teacfi
NOTE-: -Starting Solid Composition (.): 17.116*.20 Cr. 0.3610.07 Fe.
0.1410.10 Cu. 0.6910.13 Hi.
0.0610.03 Zn.< O.L. Cd.
0.3510.03 Al
'All roasts conducted in open ••rucibles in a box furnace. NaCH
added as a solution (500 gp') tu dry sludge powder.
•.All roasted soliJs were leached: 10 w/o solids, 0.5 Hr.. 25°C, pH - 7.
387
-------
If the chromium has not been oxidized then a conceptual modification of
the flowsheet presented in Figure 6.1 is that after Fe. Cu, Zn removal, nickel
can be removed by sulfide precipitation. A design matrix illustrating sue1) a
separation is presented in Table 8.111. Further testworn has not been
conducted but the approach shows a possible alternative treatment procedure.
It would mean that the resulting solution would have to be further treated to
recover the Cr* .
After chromium oxidation, as described in sections 6.3.6 and 8.9.1.1. the
solution contains only Cr (as CrJOj") and Ni with only small residua"
concentrations of other metal ions. Chromium (+6) can be separated from the
leach so'ution by: 1) precipitation, 2) ion exchange using an anion exchange
resin or 3} by solvent extraction.
8.10.1. Lead Chromate Precipitation
Chromium can be selectively precipitated from a sulfate solution by lead
cations, i.e., the Cr concentration in equilibrium with lead chromate at a pH
« 3.1 is 0.87 mg/1. Therefore, lead cations should strip Cr* from an acid
solution at a pH value near the solution pH resulting from the previous
oxidation unit operation. The chromium stripping process is illustrated in
Figure 8.24.
Small scale testwork was conducted to detemire the extent of anount of
PbS04 added. The results are presented in Table 8.112 and show that effective
chromium removal is achieved at pH values in the range 4-5.
The results of large scale lead chromate precipitation are presented in
Table 8.113. The data show that chromium can be effectively stripped from
solution, i.e.. chromium levels of 8 mg/liter were achieved.
8.10.2. Chromium Solvent Extraction
Solvent extraction literature* ' sho-s that Cr* as the anions, chromate
or dichromate, can be selectively extracted from acidic sulfate solutions by
Alamine 336 (a Tertiary Ami re) and Aliquot 336. The extraction is pH
388
-------
PH . 3-4
Ni Bearing
Solution
H CrO Solution
Figure 8.24. Chromium Removal by Lead Precipitation
389
-------
TABLE 8.1M. DESIGN MATRIX FOR SELECTIVELY REMOVING NICKEL FROM CHROMIUM (CR***J SOLUTION (FULL REPLICA)
s
a
•
•
2020
?0302t
Bin
Unit
High |.)
Ion (-)
leit f
1
2
]
4
S
6
T
a
But
Effects (I)
N1
Cr
Solution
PH
2.5
1
3.5
1.5
-
*
-
*
-
*
-
*
32.7
2.3
!!••
loin.)
15
5
20
10
-
-
t
»
-
.
*
»
-2.6
-0.5
Aiojnl
• *.S
(Sloich.l
1.5
.5
2.5
1.0
-
-
-
-
*
»
t
4
5.0
0
Results: Extraction fron Solution (X)
"wl
22.6
89. ?
-------
TABLE 8.112. CHROMIUM REMOVAL BY LEAD CHROMATE PRECIPITATION
Sample
2197
2289
2291
2215
2297
2299
2226
2301
2303
2206
2237
2305
2307
2311
No. Conditions Chromium
Starting Solution (0.31 gpl Cr)
Oxidized with C12 fof one hour
at pH • 1.5.
Contacted with 5 g PbSOa
Contacted with 10 g PbS34
Starting Solution (0.31 gpl Cr)
Oxidized wth C12 for one fnur
at pH • 3.7.
Contacted with 5 g PbSO.
Contacted with 10 g "
Starting Solution (0.31 gpl Cr)
Oxidized with CU for one hour
at pH • 3.9.
Contacted with 5 g PbSO.
Contacted with 10 g — *
Starting Solution (2.61 gpl Cr)
Oxidized with Cl. for one hour
at pH • 1.3.
Contacted with 5 g PbSO.
Contacted with 10 g "
Starting Solution (2.61 gpl Cr)
Oxidized with Cl« for one hour
at pH • 4.0
Contacted with 5 g PbSO,
Contacted with 10 g *
Solution 2307 readjusted
to pH • 5. Contacted with
5 additional grans of PbSO^.
Precipitated (*}
0
0
65.5 (pH decreased
62.9 (
86.6
89.8
0
0
to 3.4)
•
95.6 (pH decreased to Z.-8)
95.2 ( ' )
88.3 additional removal
Notes: 100 cc exposed to chlorine; pH maintained during oxldatlcn; pH
adjusted to desired level; contacted wltn PbS04 for 10 minutes.
392
-------
TABLE 8.113. LARGE SCALE TESTUORK ON LEAD CKROHATE PRECIPITATION
I
Saaple No. Condition Concentration Igpl)
Cu Fe Zn Cr Ni Cd A1
42 liter Test
2600 Starting Solution. pH • 4.2. 0.03 0.03 0.06 2.34 I.57 0.10 0.16
2601 Five Bin. exposure 0.03
-------
Independent and can be applied over the range pH 1-7. The disadvantage is
that the amine reagents are degraded by acidic solutions over long periods of
time<5>.
8.10.3. Cichromate Ion Exchange
Anionic ion exchange (IX) is a means of selectively extracting chromium
anions from chromium-nickel sulfate solutions, e.g.. Rohm and Haas IRA 900 (a
strongly basic IX resin) will quantitatively remove chromium anions from Cr-Ni
solutions. However, recovery of the chromium from the resin by stripping with
NaOH is difficult. Stripping can be accomplished but large volumes of strip
solution are required. Therefore, the concentration of the recovered chromium
is rather low. This is a distinct disadvantage because weak chromium bearing
solutions require concentration by solution evaporation which adds cost to the
overall recovery process.
An additional potential problem with IX resins is degradation oy the leach
solution, especially highly oxidizing solutions such as high concentration
dichromate bearing solutions. Repeated load/strip testwork in the present
study definitely released the characteristic ammonia odor to the laboratory.
Chromate recovery from plating rinse waters is commercially practiced on
solutions of low chromium content (<100 mg/1) and at pH values in the range of
4.5-5.0 using a weakly basic anion exchange resin, e.g., IRA 94. Such
solutions are drastically different from those solutions considered in this
study, e.g., tne present leach solutions are highly oxidizing and contain much
higher chromium concentrations.
8.11. NICKEL EXTRACTION
The nickel concentration In the treated large scale leach solution is in
the range of a few grams per liter. Leach tests on segregated sludge materials
produced nickel contents up to 42 gpl. However, most mixed metal sludges
produced nickel contents in the range 2-6 gpl. A means of recovering the
nickel from solution and/or concentrating it in solution is required.
394
-------
8.11.1. Sulflde Prec1p1t3t1on
Nickel sulfide can be effectively precipitated from a nickel bearing
solution by the addition of a sodium sulfide solution. If the sulfide solution
Is added at tne proper concentration and rate there is no release of hydrogen
sulfide gas. The pH of the leach solution following lead chromate
precipitation is 4-4.5. It Is desirable to maintain the pH at thi's level (or
even higher is better) to prevent the release of hydrogen sulfide gas.
The results of sulfide precipitation from the large scale tests are
presented in Table 8.114. The nickel content was decreased to six mg/liter,
the chromium content to four rag/1 Her, and all other metal values to below
their detection limit.
Realistically tne amount of sulfide added to precipitate the nickel would
be chosen to less than the nickel stoichiometric requirement. It would not be
a problem if some of the nickel were left in solution since most (>90%) of the
final solution Is needed as make-up water in the leach unit operation. A
deficiency of sulfide would be required because if sulfide existed in the
make-up water then H_S would be produced in the leach stage.
An alternative treatment approach would be to precipitate nickel hydroxide
along with the lead chromate by raising the pH to the range 6-9. The residue
could then be releached in ammonium carbonate to redissolve the nickel
hydroxide as a nickel amine. The lead chromate wou'd not be dissolved.
The advantages of this approach are two-fold. First, the filtrate from
the lead chromate-nickel hydroxide precipitate solid/liquid separation could be
recycled without fear of H-S generation in the leach stage and the ammonium
leaching of the residue would produce a concentr/-ced nickel solution that could
be treated to produce nickel sulfate, nickel carbonate, or other nickel
compounds.
8.11.2. Solvent Extraction
Commercial solvent extraction of nickel from sulfate solutions is not
extensively practiced. The equilibrium distribution diagrams show that nickel
395
-------
OJ
VO
o»
TA3LE 8.114. LARGE SCALE
Sample No. Condition
2605
2606
260?
2609
2610
Starting Solution,
sequentially treated test
series five, pH •• 4
Ten nin. exposure
Twenty min. exposure
Forty-five nin. exposure
Sixty nin. exposure
TESTWORK ON NICKEL SULFIDE PRECIPITATION
Concentration (gpl)
Cu __F_e__. ?n Cr Ni
0.04
-------
is not extracted from su-lfate solution at low pH levels. I.e., refer to Figures
8.10a., 8.10t>., 8.12., 8.16 for O^HPA, YE3SATIC Acid 911. LIX-64N. The pH
level required for nickel extraction is too high to be applied to a chromium
bearing solution, i.e., Cr* will begin to precipitate at pH levels >2.5-3.5
(depenas ot. concentration in the solution). Therefore, the solvent extraction
of nickel as a means of separating Ni from Cr is not feasible. Solvent
+2
extraction of Ni from the final leach solution, efter Fe, Cu, Zn, ana O are
removed, appears to be possible by O.EHPA. If the extraction is possible then
SX would be a way of concentrating the NI content into a strip solution and NiS
precipitation would be unnecessary. Bench scale shake tests were performed and
the results are reported in Tables 8.115.-8.117.
The Influence of pH on LIX-64N extraction of N1 was investigated (Table
8.115); about 20 percent extraction was achieved at an initial pH =» 4 and pH *
5 in two contacts. At an initial pH of 9 (Table 8.116.) 86 percent was
extracted. The use of an NH4OH, HH4C03 buffer solution gives almost
quantitative extraction of Ni (Table 8.117.). However, the problem with the
high pH systems 1s that a portion of the nickel is precipitated from solution
as a hydroxide.
Although other investigators have shown extraction of nickel by DgEHPA at
pH values of 5-6, the present test work did not. The results of a series of
shake tests are presented in Table 8.118.
Recent developments^ ' in SX show that a pronounced synergistlc effect
occurs when D.EHPA is mixed with non-chelating aldoximes. The extraction
sequence is drastically altered by the presence of the non-chelating aldoxime:
order of extraction by D-EHPA:
Zii>Cu>Fe>Co>Ni
Order of extraction by D.EHPA plus 2-ethylhexanol oxime (EHO):
N1>Cu>Co>Fe>Zn
397
-------
TABLE 8.115. LIX 64N EXTRACTION Of NICKEL AS A FUNCTION OF PH: PH « 4-6.6
Sanple No. Condition
1370 Potassium Jaroslte
Concentration (gpl)
Fe Cu Nl __Cr_ Zn_ Cd Al
0.42 3.69 2.10 0.54 6.89 0.27 4.65
CO
In-sltu prccipltatl on ;
stx-hour exposure
Copper Removal
5olutloini75~pll adjusted
to 1.75. Contacted twice
with 20X LIX 622
1372
1373
1374
1375
1376
1377
First Contoct
Second Contact
Zinc Removal
Solution 1373 pH adjusted
to 2.0. Contacted four
times with 40S OEMPA
First Contact
Second Contact
Third Contact
Fourth Contact
0.48
0.46
0.38
0.024
0.013
0.007
0.34
0.010
0.009
0.010
0.009
0.010
2.10
2.21
1.73
2.02
1.74
2.20
0.58
0.61
0.47
0.56
0.47
0.60
7.22
7.65
2.23
0.83
0.19
0.02
0.30
0.31
0.24
0.28
0.23
0.26
4.77
5.36
3.58
3.06
1.53
0.64
-------
TABLE 8. IIS. CONTINUED
Sample No. Condition Concentration (gpl)
Fe Cu Ni Cr In Cd Al
Adjustment (NHaOH) and
PH Adjustment (NHaOH) a
L1X 64H (IQt) contactr
1385 Solution 1377 after pH 0.003 0.002 1.22 0.28 0.00 0.14 0.11
adjusted to 4.0
1388 Aqueous after first
-------
TABLE 8. MS. CONTINUED
Sanple Ho.
1395
1398
1387
1390
1393
1396
1399
Condition
First Strip of 1392
Second strip of 1392
pH !_•_ 6.J>
Solution 1377 after
pH adjusted to 6.6
Aqueous after first
contact (0/A • 1)
Aqueous after second
contact
First strip of 1393
Second strip of 1393
Concentration (col)
Fe
-------
TAOLE 8.116. LIX 64N SOLVENT EXTRACTION OF NICKEL: NH^OH, NH4C03
Sample No. Tondltlon Concentration (gpl)
Fe Cu Hi Cr In Cd Al
1377 Starting solution: 0.007 0.010 2.20 0.60 0.02 0.26 0.64
potassium jaroslte solu-
tion treated for Cu, Fe,
2n removal
1«33 Solution 1377 (ISO cc)
-------
TABLE 8.116. CONTINUED
Sample Ho. Condition Concentration (gpl)
Fe Cu HI Cr in Cd Al
1439 Strip of 1436 organic with
-------
TABLE 8.117. BENCH SCALE SEQUENTIAL SOLVENT EXTRACTION TESTMORK
Sample No. Condition ' Concentration (gnl)
Cu Fe Zn Cr HI Cd Al
1466 Jarostted leach solution; 3.14 1.44 9.37 O.S4 4.95 O.S2 1.58
one liter
1467 Diluted 1466. pH • 1.75 1.56 0.69 4.99 0.27 2.75 0.24 0.63
Cu SX
1468 LIX 622 (lOv/o) contacted with 0.02 0.69 5.00 0.27 2.74 0.25 0.64
- 1467 (0/A - 1)
o 1469 LIX 622 contacted with aqueous 0.004 0.68 4.93 0.?6 2.71 0.25 O.C3
M from 1468. Initial pll of aqueous
to 1.75
Zn SX
1470 Aqueous 1469 adjusted to pH 0.001 0.63 1.71 0.27 2.76 0.23 0.55
• 2; then contacted (first
contact) with DEHPA organic
(0/A -1)
1475 Aqueous 1470 adjusted to pH < D.L. fl.57 O.32 0.27 2.74 0.19 0.39
• 2; then recontacted (second
contact) with uEIIPA.
1477 Aqueous 1475 adjusted to pH < D.L. 0.44 0.03 0.25 2.59 0.12 0.16
• 2; then recontacted (third
contact) with DEHPA.
-------
TABLE 8. 117. CONTINUED
Sample
1479
1471
1473
1472
1474
1476
1478
1480
No. Condition
Cu
Aqueous 1477 adjusted to pll < D.L.
-2; then rccontacted (fourth
contact) with DEIIPA.
DEHPA Organic Strip (Compositions of
Organic 1470 stripped with < 0.001
200 gpl sulfurlc acid
Above organic stripped second "
time with 200
-------
s
Ul
TABLE 8.117. CONTINUED
Sanple
14C2
1484
1485
1486
1487
No. Condition
Cu
Aqueous 1481 contacted with < D.L.
LIX 64N (0/A • 1)
Organic 1482 stripped with < 0.001
200 gpl sulfurlc acid
N1 SX: LIX 64N (lOv/o)
PH raised to 9 with < 0.001
tU^OH/fOMCOj mixture (40 gpl
each), filtered
Aqueous 1485 contacted with < D.L.
LIX 64N (0/A - 1)
Organic 1486 stripped with < 0.001
200 gpl sulfurlc acid
Concentration (gpl)
Fe Zn Cr
3.15 < D.L. 0.13
•
0.02 < DA. < 0.01
0.02 < D.L. 0.02
0.02
c 0.001
Ni
0.15
1.30
0.48
0.08
0.35
Cd Al
0.06 0.02
< D.L. < 0.01
« O.L. < 0.01
• *
" < 0.001
Notes: . Barrel 14 sludge leached undar standard conditions.
. Cu SX: lOOcc of aqueous contacted with lOOcc of organic. 3 nln. anfcient temperature,
0/A - 1
. Zn SX: lOOcc of aqueous contacted with lOOcc of organic. 40 v/o DEHPA. conditions as
above.
• DEHPA Strip; conditions as for other SX tests.
. N< Sidconditions as above, pH of aqueous phase raised, solution filtered, then contacted
with organic.
-------
en
TABLE 8. MB. DEIIPA EXTRACTION
Sample No. Condition
1377 Starting solution:
potassium Jarosite solu-
tion treated for Cu, Fe.
Zn removal. Pll '1.29
1458 Solution pll adjusted to
to 5.0 and filtered
1459 Solution pll adjusts to
6.0 and filtered
10 v/o OEHPA - 90 v/o
47UB; pH • 5.0
1442 First contact with 1458
Final pll - 2.6. 0/A - 1,
R.T.. 2 minutes
1443 Strip of 1442 organic
with 150 gpl H.SO.,
0/A - 1 z 4
1444 Second contact, aqueous
OF rilCKEL AS A FUNCTION OF SOLUTION Pll: PH « 4.5
Concentration
Nl Fe ' _CJL_ _ Cr
2.20 0.01 0.01 0.60
1.64 0.21
O.BB
-------
TABLE 8.118. CONTINUED
Sample No.
1450
1451
1452
1453
1446
1447
1448
Condition
Nl
pH • 6.0
First contact of 10 v/o 0.87
DEHPA with 1459 aqueous.
pH- 6.0
Strip of 1450 organic 0.01
«tth 150 gpl H2S04
Second contact, aqueous 0.77
1459. pH • 6.0
Strip of 1452 organic 0.03
with 150 gpl H2S04
40 v/o DEHPA. 60 v/o 4708
pH - 5.0
First contact, aqueous 1.57
(1450) with 40 v/o DEHPA.
0/A • 1
Strip of 1446 organic 0.08
with 150 gpl H2S04
Second contact, aqueous 1.58
(1458) with 40 v/o DEHPA
Concentration (gpl)
Fe Cu Cr Zn Cd Al
0.01
-------
TABLE 8.118. CONTINUED
Sample NJ.
1449
1454
• 1455
1456
1457
Condition
Strip of 1448 organic
with 150 gpl H2S04
First contact, aqueous
(1459) with 40 v/o DEHPA
Strip of 1454 organic
with 150 gpl H2S04
Second contacv, aqueous
(1459) with 40 v/o OEMPA
Strip of 1456 organic
with 150 'jpl H2S04
Concentration (gpl)
N1
0.10
0.80
0.02
0.84
0.06
Fe
< D.L.
0.01
0.01
0.01
0.01
Cu
0.01
< D.L.
< D.L.
< D.L.
« D.L.
Cr
0.04
0.02
0.01
0.02
0.01
Zj>
< D.L.
< D.L.
< D.L.
< D.L.
€ U.I.
Cd _AJ
0.05 < D.L.
< D.L. < D.I.
0.08 < D.L.
< D.L. « D.L.
0.01 < D.L.
Notes: . Starting solution treated for Fe. Cu. Zn removal before nickel testwork.
. Contact conditions: 0/A • 1. Temperature • 25°C, time • 2-3 rat mites.
-------
TABLE 8.
Saonle No.
5056
5057
SOS8
50S9
5060
119. NICKEL
Final pH
3.1
3.0
?.l
3.0
3.0
EXTRACTION BY 40 V/O
DEIIPA - 70 GPL
Nt Concentration In Aqueous
Phase (gpl)
Initial
4.37
4.26
5.10
S.OS
S.3I
Final
2.97
3.30
3.84
3.30
2.92
ENO
Extraction
32.0
22.5
:4.7
34.6
45.0
Average 31 .6
(gpl) Strip (gpl
100.0
100.0
100.0
100.0
100.0
Notes: . 0/A • 1 loading. 0/A • I stripping. SO cc each phase. 25°C
. 200 gpl sulfuric acid
-------
The change in order is not as important to the present project as the fact that
the pH at which extraction occurs is shifted dramatically^ ', Figure 8.2?,
i.e., the pH for 50% Ni extraction by D2EHPA-SHO is 1.58 while the pH for 50%
Ni extraction by D^EHPA alone is 4.11. The DgEHPA-EHO mixture appears to be
worth further consideration as a means of concentrating the Ni content.
Figure 8.25. Extraction of nickel, calcium, magnerium by 0.5 M D-EHPA
and its mixtup«,with 0.5 M 2-ethylhexanol oxime (EHO)
(from Preston**•*').
A series of small scale shake tests were performed to verify that nickel
could, indeed, be extracted from low pH solutions by a O^EKPA-EHO mixture and a
LIX63-D-EHPA mixture. Also it was important to determine the selectivity of
the organic for nickel in the presence of chromium.
A series of shake tests was performed at a pH of approximately three using
f44}
the organic composition suggested in the study by Prestonv '. The results
shown an average extraction of approximately 32 percent nickel, Table 8.119.
A design matrix series was run and the results are reported in Table
8.120. Excellent extractions were achieved for several conditions, e.g.,
80.2-83.1%, for a single contact. The stripping is also excellent. This
system is far from being optimized. A graduate student is continuing the
research beyond the results reported here. The system does appear to hold good
potential for nickel recovery and concentration. If nickel could be solvent
extracted from the chromium (+3) at a pH <2.5, then the expensive chromium
410
-------
TABLE 8.120. DESIGN MATRIX FOR OEHPA - EHO SOLVENT EXTRACTION SYSTEM
Sample No.
5061
S062
5063
5064
5055
5066
5067
5068
5069
5070
5071
Tine
(Bin)
3
7
3
7
3
7
3
7
7
5
5
£110
<9P»
70
7fr
130
130
70
70
130
130
130
100
100
DEHPA pH(Ftnal)
-------
oxidation would be unnecessary. The nickel could be extracted and recovered by
electrowinning and the chromium could be precipitated as Cr(OH)-; then calcined
to CrpO,. This system is presented as one of the cost alternatives in Section
8. IS. A considerable saving in cost may be possible.
Another system has also been examined in a cursory manner; the LIXoS-
D-EIPA system. An isotherm was run at a pH of 1.17 using a mixture of 12.5 v/o
LIX63, 16.1 v/o OgEHPA, remainder KERMAC 510. Good extraction is Indicated by
the data presented in Table 8.121. Chromium is not extracted. Testwork on
both the above systems is continuing. Several variables will be investigated:
the organic composition (EHO or LIX63 mixed with D-EHPA) ; equilibrium pH; time
of contact; temperature; stripping efficiency as a function of acid strength.
Both systems appear t£ ho}£ promise for nickel extraction a£ low pjl levels.
8.12. FINAL LARGE SCALE SOLUTION PURIFICATION
A sample of the large scale test (sequential test series five) leach
solution (after Fe, Cu. Zn, Cr and Ni -emoval) was treated by a cation exchange
resin. The solution after nickel sulfide precipitation contained only 4 mg/1
Cr and 6 mg/1 Ni . All other elements were less than the analytical detection
limit (<0.01 mg/1). A 150 cc sample was contacted with IRA-200, a strong
cationic exchange resin in the hydrogen form and a second sample was contacted
with the same type resin in the sodium form. The final Cr and NI content was
reduced to < 1 mg/1 and 1 mg/1. respectively, using the H form; and to 3 mg/1
and 2 mg/1 respectively using the Na form.
Practically all of the final solution will be recycled to the leach unit
operation, i.e., for the treatment of 100 pounds of sludge, most of the final
solution volume will be recycled to satisfy the leach water requirement.
Therefore, if IX is required as a final clean-up unit operation, it will be
necessary to only treat a rather small volume of solution.
8.13. LARGE SCALE SEQUENTIAL METAL EXTRACTION AND RECOVERY TEST DATA
A series of large-scale tests were conducted to develop test data to
design a particular unit operation; to design further testwork; or to verify
4i2
-------
TABLE 8.121. NICKEL
Sample
5071
5072
5073
5074
5075
5076
5077
Notes:
No. Condition
Starting
0/A • 10/1
5/1
2/1
I/I
1/5
1/10
. Solution pH 1
. Organic: 12.5
SOLVENT EXTRACTION BY LIX63
-DEHPA: ISOTHERM DATA
SUMMARY
Concentration (gpl)
HI
5.544
0.695
7.038
3.325
3.685
4.368
5.059
.17
v/o LIX 63,
Cr
10.525
10.728
10.918
10.669
10.604
10.529
10.457
16.1 v/o
CIL
0.022
0.007
0.011
0.040
0.022
0.013
0.014
DEHPA,
Je^
0.007
0.005
0.007
0.007
0.007
0.008
0.008
renal nder
Jn
0.020
0.004
0.008
0.007
0.017
0.014
0.015
KERHAC 510
Pb
0.002
0.013
0.005
0.022
0.018
0.010
0.005
Temperature: 2SOC
. Tine: 3 minutes
-------
the applicability of a particular unit operation. These data are presented in
the following tables and are referred to and discussed throughout the previous
sections. The data are in a chronological order as collected. Not all unit
operations were performed in every sequential test.
8.13.1. Sequential Test: Series One (83 pound Test)
Purpose of test: Generate solutions for copper and zinc solvent
extraction.
Results
Comments
Table 8.122. Sequential Test:
Table 8.123. Sequential Test:
Series One
Non-Recoverable Elements
First use of filter press resulted in the discovery that
the diaphragm pump was not stainless steel as per
specification. Pump contaminated filtered solution with
iron. Test was not desicned to collect mas; balance data
but was performed to gain experience with the large
scale leach, filter press and Sx equipment. Phase
separations in the SX testworic were good, flpwrate of
solutions controllable, muck and crud formation minimal.
8.13.2. Sequential Test: Series Two (200 Pound Test)
Purpose of Test: A large leach was conducted then split into two volumes.
One volume was saved to be used as stock solution for
feeding into operating jarosite solution. The test
objective was to investigate continuous jarosite
precipitation under constant feed conditions. A second
purpose was to prepare a large volume for SX work.
Results
Comments
Tables and 8.124, Sequential Test: Series Two
Table 8.123, Sequential Test: Non-Recoverable Elements
Continuous precipitation test not run because desired
iron level in operating jarosite was not achieved. Zn
SX did not remove all zinc in first series of contacts
because insufficient D-EHPA present. Zinc raffii.ate
recycled through system twice more at higher starting pH
levels. Gypsum precipitated in the strip cells.
8.13.3. Sequential Test; Series Three
Purpose of Test: Further large scale testwork on copper, zinc SX.
Results : Table 8.125, Sequential Test: Series Three
Table 8.123, Sequential Test: Non-Recoverable Elements
414
-------
Ul
Saiaple No.
1366
1367
1371
1440
1466
. TABLE 8.122. SEQUENTIAL TESTS:
Conditions
_fe_
Leach- Jarosite Precipitation
Leach solution. 53 C; 20.20
(one-half hour, standard t '
conditions); 64 liters [Fe * -
Leach solution, diluted and 14.00
pH adjusted to 2.14 . 90 liters
SERIES ONE (83 POUND TEST)
Cu
5.15
0.91 gpl
3.60
Concentration (gpl)
JJL Cr 7n Cd Al
8.04 1.08 13.96 0.77 2.23
5.45 0.75 9.89 0.53 1.56
V
five hour exposure to 4.02 3.62 6.32 0.65 11.38 0.62 1.77 (Vol. -78
potassium Jarosite conditions, t .. lit.)
In-situ deposition of Jarosite [Fe ] • 0.92 gpl .
Ifter'fiPeHna^hrouMlter Convositlon adjusted for solution loss frm vessel
arier filtering tnrougn nicer * .7 ... ... n ., _ ftl ft -, ...
press. pH • 1.03 final. 90 3'*T 3tlz 5>" °'56 9'81 °-53 '-53
liters of solution (later
found to be Iron contaminated
by filter press pump)
Abo-.e solution (1371) p!< 3.36
adjusted with K Oil to 2.75. volume 99
Seven-hour exposure to 1.57
potassium Jarosite
3.14
liters.
2.87
4.95 0.54 9.37 0.52 1.58
5.04 0.58 9.86 0.52 1.30 (Vol.. 94
lit.)
deposition. Filtered one- Composition adjusted for solution loss fron vessel
balf batch before punp 1.49 2.72 4.79 0.55 9.36 0.49 1.23
failed. pH final - 1.90
(see Note)
-------
Ol
TABLE 8.1
Saiiple No. Conditions
-It-
Copper SX (10 v/o LIX 622); Relster
0/A • 1. 2-stage extraction;
1 -stage strip (150 gpl H2S04.)
recycled; 250 cc/mtn. flow
rate
1523 Solution 1466 diluted to 0.65
•40 liters. Starting solu-
tion for Cu SX test (40
liters) Initial pH • 2.14
1524 Rafflnate from contact. 0.68
Final pH • 1.73. 40 liters
1526 Final strip solution; 3.8 1; 0.028
(starting acid 150 gpl H.SOJ;
Note some solution carry over.
Zinc SX (27 v/o OEMPA
0/A • 1, 4-stage extraction;
Z -stage strip (200 gpl 112804)
recycled; 250 cc/mln. flow
rate., Initial pll « 1.75. 40 liters
1532 Raftlnate from Contact. 0.46
final pM • 1.29. 25 liters
1533 Final strip solution, 7.6 1; 0.005
22. CUiUINUED
Concentration (gpl)
Cu .HI Cr Zn Cd
Systeia;
1.37 3.02 0.28 4.94 0.31
0.017 3.18 0.29 5.14 0.33
•
.7.22 0.10 0.018 0.076
-------
TABLE 8.122. CONTINUED
NOTE: 'Barrel 4 sludge material. Sludge weight: 83.1 pounds. Solids weight: 27.4 pounds.
•Sljdge Composition (1): Fe (14.2). Cu (3.73), Nl (4.60). Zn (8.87). Cr (0.72). Cd (0.48).
Al (1.58).
•The filter press punip contaminated solution with Iron. The Iron content of solution before
filtering was 0.068 gpl (68 ppm). After filtration the Iron content was 1.44 gpl (sample 1466).
.MX 622 loads copper that docs not strip completely. This effect Is characteristic of LIX.
.DEMPA concentration Insufficient to remove all the zinc; DEIlPA loids 0.15 gpl Zn per v/o
DEHPA; therefore. 27 v/o DEItPA will extract only 4.0 gpl Zn. Also, later found that pH
decreased in first two contacts to 1.3; therefore, last two stages of contact were not extracting
Zn into the organic phase.
.Jaroslte conditions: pH: 2-2.8, temperature: 88-92.
-------
TABL; 8.123. SEQUENTIAL
TEST:
HON-RECOVERABIC ELEMENT?
Concentration (spl)
Sam>le
1366
1367
1371
1466
1523
1524
1526
1532
1533
No. Condition
SERIES ONE (For con
Leach Solution
Diluted
5 hr Jaroslte
7 hr
Start for Cu SX
Cu Raff mate
(Also start for
Zn SX)
Cu Final Strip
Zn Rafflnate
Zn Final Strip
SI
ditlon
1.56
1.09
1.16
0.12
0.12
0.03
0.11
0.01
SERIES TUO (For condition
1765
1757
1769
1802
1816
1817
1811
1824
1825
1826
1833
1834
1882
1889
1890
Leach Solution
1 hr Jaroslte
Final * (12 hr)
Cu SX Feed
Cu Rafflnate
Cu Final Strip
Zn Fee4(pH*1.75)
Zn fatflnate.3.25
hrs.
Zn Strip
Zn Feed(pH-2.02)
Zn Rattlnate.2.7S
hrs.
Zn Strip
Zn Feed(pH-2.5)
Zn Rafflnate. 3 hr
Zn Strip
2.16
1.78
0.19
0.28
0.28
0.01
0.28
0.27
0.03
0.27
0.27
0.08
0.22
0.22
0.03
Ca
details
0.58
0.53
0.60
0.51
0.30
0.31
0.06
0.17
0.13
details
0.12
0.13
0.05
OJ7
0.05
0.42
0.02
0.21
0.13
2.45
0.02
0.18
0.26
Al
see Table
2.23
1.56
1.77
1.35
0.55
0.56
0.03
0.32
0.03
see Table
2.27
1.87
0.56
0.44
0.45
0.03
0.45
0.28
0.17
0.29
0.06
0.52
0.05
O.L.
0.65
Fe
8.86)
20.20
14.00
4.02
1.44
0.65
0.68
0.03
0.46
0.01
8.88)
16.33
2.87
1.04
1.14
0.11
1.13
1.00
O.L.
0.97
0.60
« O.L.
0.52
0.16
0.03
p
3.09
2.04
1.48
0.78
0.69
0.16
3.17
0.17
-------
TABLE 8.123. CONTINUED
Sample
1962
1966
1991
2005
2066
200S
2109
2110
2116
2118
2126
2127
2144
2143
2146
2147
No. Condition Concentration (c?l)
SI Ca
A1
SERIES THREE (For condition details see
Final Jaroslte 1.1S 0.80 2.99
solution "~
Large System Cu 1.16 2.48
S!F; rina'i raff mate
after crud shutdown
Snail System Cu 0.68 0.54
5X; leach solution
BTluted. starting solution.
Cu Final Raff. 0.63 0.45
Cu Final Strip 0.02
-------
TABLE 8.123. CONTINUED
Sample No. Condition
2177
2131
21 BOB
2242
2256
2181 a
2361
2340
2374
2376
2348
2352
2364
2367
2378
2492
2494
SI
Zn SX Starting 0.88
Solution. pH • 2.0
Final Raff. 2.5 0.86
hrs.
Final Strip 0.02
Zn SX Starting 0.16
Solution, pH • 3.0
Final Raff. 6 hr 0.48
Chromium 0x1 d. 0.86
Starting solution
before pH adjust
pH adjusted to 0.07
5. aqueous phase
Chlorine oxld. 0.07
1 hr
Chromium Reoxid. 0.03
I.b HI*.
Chromium Predp. 0.03
by lead suifate
added to 2374
Chromium Predp. 0.05
7f, starting solution
One hr exposure. 2X
PbSO. added 0.05
Lead'Sulfate- «D.L.
Lead Chromate
Nickel Preclo. 0.04
froa 10 liters
of 2352. starting
solution
Final Filtrate 0.03
SERIES FIVE (For condition
Leach Solution 0.70
Final Jarosite 0.87
Solution
Concentration
Ca
0.40
0.07
0.46
0.32
0.08
0.07
0.08
0.10
C.07
0.07
0.09
0.09
0.10
0.04
0.04
Al
1.22
0.49
0.34
0.81
0.32
0.49
-------
Sample
2494
2499
26008
Z52S
2S26A
2S24
2S26B
2527
2564
2592
2600
2603
2605
2610
TABLE 8.123.
No. Condition
SI
Cu SX Starting 0.87
Solution
Final Raff. 0.82
Final Strip O.I.
Zn SX Starting 0.82
Solution, pH • 2.0
Final Raff.for 0.79
starting pH • 2.0
Final Raff, for 0.51
starting pH • 2.5
Composite Raff. 0.75
Final Strip 0.09
Chromium Oxld. 0.86
starting solution
Final 0*1 d. 0.24
solution, 83.61 oxidized
Cnromlua Preclp. 0.27
Final Exposed 0.17
Solution, 45 Bin.
N1S Precipitation 0.23
starting solution
Final exposed 0.19
CONTINUED
Concentration (gpl)
Ca
0.21
0.22
0.03
0.22
0.07
0.04
0.04
0.49
0.04
0.05
0.04
0.04
0.03
0.03
Al
1.30
1.25
O.L.
1.20
0.40
0.52
0.38
8.27
0.39
0.08
0.16
0.14
<0.01
-------
TABLE 8.124. SEQUENTIAL TEST: SERIES TWO (200 POUND TEST)
IN)
Sample No.
1765
I7S7
1769
1802
1802
1797
1799
1805
1809
1DIC
Condi tl on
leach-Jarosite Precipitation
Starting Solution, 'one-half
hour standard leach; about
one half of leach solution
subjected to jaroslte
conditions, 60 liters
One hour exposure
Twelve hour exposure.
pll overshot drastically
Final solution after
filtration. pH adjusted
to 1.9.
Copper SX
Feed Solution as above
r«rrr Raffinate (Starting
1/2 hour
1 hour
2 hours
3 hours
t hours
Concentration (gpl)
Fe
•MM^^W •
16.33
2.87
1.04
1.13
pH-1.9)
.12
.13
.14
.16
.14
Cu
1.48
1.08
1.04
0.39
0.003
0.01
0.06!
6.07
•EBfi
Zn
9.96
8.57
8.31
8.89
' 8.89
' 8.89
t 8.89
8.91
> 8.84
Cr
•MM*^ •
1.05
0.69
0.55
0.26
0.26
0.26
0.26
0.27
0.27
HI
•^H^MW •
8.47
7.12
6.94
8.02
7.87
7.99
8.00
8.22
8.07
Cd
0.46
0.39
0.38
0.41
0.41
0.41
0.41
0.43
0.42
Al
2.27
1.87
1.68
0.44
0.45
0.45
0.45
0.46
0.45
-------
TAHLE R.124. CONTINUED
Sample No.
Condition
Zn SX (Feed from Cu SX
adjusted to pll - 1.75)
Concentration (gpl)
Fe
Cu
Zn
Cr
Nl
Cd
Al
ro
1811
IBM
1818
1824
1826
1827
1831
1833
1882
1885
1889
NOTE:
Feed Composition
l.S hours
2.5 hours
3.25 hours
Above Raffinate Adjusted to
pH - 2.02 and Recycled through
System
pH Adjusted Feed
1 hour
2 hours
2.75 hours
Above Raffinate Adjusted to
pH • 2.50 and Recycled through
System
Feed
1 hour
3 hours
1.13
0.97
0.93
1.00
0.97
0.60
0.59
0.60
0.60
0.12
0.058
0.051
0.034
O. L.
0.005
<0. L.
0.011
0.013
6.013
0.002
Starting SljJge Composition (average of 7 samples In
Barrel 14 sludge; 81 .6t H?0; 18.61 solids.
.
8.84.
1~67~
3m
1.63
EOT
DTK
O5
«-ptf!
X): 18.69
1.07
2.25
0.27
0.27
0.25
0.27
0.27
0.26
0.26
0.27
0.27
0.25
0.25
» 1.35
* 0.03
7 0.28
8.02
8.06
7.65
8
8
7
e
8
8
8
a
Fe.
Cr.
Al.
.08
.10
.99
.45
.67
.67
.75
.17
10.11
8.52
0.43
0
0
0
0
0
0
0
0
0
0
0
*
7
7
.42
.40
.38
.47
.41
.29
.35
.37
.37
.17
.13
1.54
0.73
0.03
0.4b
0.27
0
0
0
0
0
0
0
0
-------
TADLC 8.124. CONTINUED
Leach-Jaroslte
. Standard leach conditions, 156 liters
. Standard Jaroslte conditions Initially established. pH overshot during las two hours of
test. Significant metal value lost by hydroxide precipitation.
Cu SX
. Two stages of contact (Initial pll « 1.90, I stage of scrub (100 gpl ^$04), 1 stage of strip
(I7C gpl feSOd, aqueous solution recycled).
. 0/A - 1. 10 v/o LIX 622. 90 v/o KEftMAC 4708.
. Organic previously usec1 and not retreated before this application.
Zn SX
Four stages of contact (0/A • ?), tm stages of strip (0/A • 2. 200 gpl H.SO. fresh solution).
40 v/o OEIIPA, 60 v/o KERIIAC 470B. f *
Strip solution recycled.
Feed rate 0.25 l/mtn for organic and aqueous.
-------
TABLE 8.125. SEQUENTIAL TESTS: SCRIES THREE (75 POUND TEST)
Sample No. Condition Concentration (gpl)
1961
1962
Fe Cu Zn Cr Nt Cd Al
Residue- Jaroslte Precipitation
Final 6 hour exposure, 82 lit. 0.50 4.07 10.55 0.72 5.61 0.58 3.07
Final solution allowed to f.46 3.99 10.24 0.70 5.36 0.56 2.69
settle 8 hours
ui
Large System Cu SX
10 v/o LIX 622. 90 v/o KERNAC 470B; two stages of K.SO. (100 gpl) scrub; two stages
of strip (150 gpl M-SO.I; aqueous phase Initial pH • 270. Systeii uncontrollable
because of crud formatton.
Snail System Cu SX
Above leach solution diluted to increase available voluoe of solution; two stages of
kerosene scrub; two stages of contact (0/A • 1); two stages nf strip (0/A • I).
1991 Startlno feed solution 0.28 2.32 5.96 0.42 3.78 0.41 1.58
2GOE Final raftlnate. approximately 0.26 0.047 5.7C 0.40 3.39 0.36 1.58
20 lit. run through system.
2006 Findl strip 0.15 28.71 0.10 0.008 0.04 < O.L. 0.03
-------
en
TABLE 8.125. CONTINUED
Sample
No.
Snail
Condition
Concentration (gpl)
System Zn SX
Fe
Cu
W^^H^VM
Four stages of contact; pH adjustment after
2005
2096
2097
2098
2100
2102
2104
2108
2109
2110
Notes:
Starting Solution
Raff.
Raff.
Strip
Raff.
Raff.
Strip
from stage
from stage
. 3 hrs
from stage
froo stage
. 4 hrs
Raff, from stage
Raff, from stage
Approximately 20
Strip. 7 hrs
. Sludge
two.
four
two.
four
two.
four
from barrels 2
3 hrs
, 3 hrs
4 hrs
. 4 hrs
7 hrs
, 7 hrs
0.26
0.08
0.03
37QT
0.08
"(ToT
joTT.
0.10
-------
TABLE 8.125. CONTINUED
Notes: . Sn&ll: Zn SX
Four stages of extraction, 0/A • 1; 40 v/o DEMPA; pH adjusted to 2.0
after first two stages; flowrate SO cc/Mtn, temperature 55 C.
Three stages of strip; 0/A • I; recycled strip acid; 200 gpl HjSOjlinitial)
-------
Comments
Leach and jarosite precipitation went very well.
Lowered ircn content from 14.49 gpl Fe to 500 ppm in six
hours. Cu SX completely uncontrolIdble; a great deal of
crud formed and the test had to be terminated. The
suspected problem was K-SO. scrub stages. SEM work
(Figure 8.26) showed large amounts of potassium and
sulfur in the crud material. Testwork performed on
small Bell system to determine problem. Problem was
overagitation in the mixer cells. This problem was
overcome by adding flo"'.ieters to each mixer cell so that
an ev«n flow was established to each cell.
Small scale zinc tests were conducted tc study pH
readjustment after two stages of contact. Results were
favorable. Gypsum formed in strip cells. A filter
system was devised.
8.13.4. Sequential Test: Series Four (35 pound test]
Purpose of Test:
Results
Comments
To produce a solution (containing
completely througn the flowsheet.
20 gpl Fc) to carry
Table 8.126. Sequential Test:
Table 8.123, Sequential Test:
Series Four
Non-Recoverable Elements
Copper SX problem described in Sequential Test three
overcome by better control of mixer agitation and
removal of K2S04 strip cells. LIX-622 content
insufficient for complete copper removal in this
continuous test so solution cycled through system again,
Thirty liters of chromium-nickel solution treated as
depicted below:
30 liters (2181)
treated by:
C12 adjusted Eh to >1000
pH maintained >4, filtered
after oxidation (2340)
15 liters treated
to PbSO. precipitation
1 hr.. pH 4, filtered
PbSO.-PBCrO.
Released in
10 v/o of
H,SO.(2364)
172 hr. filtered
Filtrate: 2367
treated with Na,S
precipitation or
NiS, pH 5, 30 minutes
I
10 liters, reoxidized
in flowing C19 for 1.5 hr.
then filtered*
1
Solids
Filtrate: 2374
treated with PbSO.
for precipitation,
I
Solid
NiS
Filtrate: 2378
428
PbSO.-PbCrO,
Solids
Filtrate:
2376
-------
TABLE 8.126. SEQUENTIAL TESTS: SERIES FOUR ( 35 POUND TEST)
N)
NO
Sample No. Condition
Cu
Fe
Concentration
Zn
Cr
(apl)
Ni
Cd
^W^MH
_A1
Standard leach (Barrel 1) . 30 liters
2114
2115
2116
2118
2119
2125
2126
2127
2129
2130
2132
2133
15 minutes
30 minutes
45 ninutes
Jaroslte
Leach Solution Dilute!. 60 lit;
1 hour Jaroslte, 57 lit.
6 hour jarosite, 51 lit.
Jaroslte settled from
solution (8 hours), 47 lit.
Cu SX (10 v/o LIX 622). 90
So • ul Ion diluted. pH - 2.01
Raffinate, 2 hours
(pH • 1.38)
Strip, 2 hours
Strip, 3.5 hours
Raffinate, 5 hours
(pH • 1.40)
11. IS
11.16
11.16
5.21
5.14
20.37
20.24
20.47
'
8.11
UT92
17.70
17.67
18.04
7.12
7.55
Adjusted composition for
4.88
5.49
0.87
OJ
7.17
8.00
Adjusted composition for
4.66
5.81
0.28
0.39
6.80
8.58
Adjusted composition for
4.55
lit.
OS
6.80
7.87
1JS
0.30
0.33
0.31
0.01
0.01
J.32
6.72
5.80
5.63
0.01
<0.01
5.72
1.75
1.74
1.76
0.70
0.53
• 7.99
8.01
7.96
3.14
3.33
solution volume
0.50
0.48
volume
0.40
0.52
volume
0.41
0.36
0.35
O.01
-------
TABLE 8.126. CONTINUED
*»
u*
o
Sample No.
2134
2142
2144
2143
2145
2146
2147
Condition
Strip. 5 hours
Final Rafflnate, 6 hrs.
Final Composite Rafflnate
Final Composite Strip
Recycled 12144) through
system
Final f.af finale. 6 hrs, 75
Final Composite Rafflnate
Final romposltc Strip. 6
hrs. (new acid at start)
Concentration
Cu
37.46
1.34
n?
42.31
l;0.109
0.116
_;!?._
'0.01
0.30
0.31
0.05
0.29
0.29
.
N1
-------
TABLE 8.126. CONTINUED
Sample No.
2242
2243
2244
2245
2246
2247
2248
2250
2251
2252
2253
2254
2255
2256
2181a
(Before
pll adjust)
Condition
Starting Solution pH
Adjusted to 3.0
Rafflnate After Second
Contact. Start
Rafftnate After Fourth
Contact, Start
Strip. Start
Rafftnate After Second
Contact. 1 hr.
Rafftnate After Fourth
Contact. 1 hr.
Strip. 1 hr.
Raffinate After Second
Contact. 2 hrs.
Rafflnate After Fourth
Contact, 2 hrs.
Strip, 2 hrs
Rafflnale After Second
Contact. 3 hrs.
Rafftnate After Fourth
Contact. 3 hrs.
Strip, 3 hrs.
Final Raffinate, 6 hrs. ;
90 liters
Chromium Slurry Oxidation
(in maintained)
30 liters of solution
(pll • 1.3. Eh • 380 ov)
2181 doped Ntth 42 gpl
(Concentration (gpl)
tii
0.11
0.11
0.09
0.02
0.11
0.10
0.02
0.11
0.10
0 01
0.12
0.11
0.04
o.io
0.08
•.
0.5B
0.39
0.02
-------
TABLE 0.126. CONTINUED
Sample Ho.
*«
to
to
2361
2340
Condition
Cr**» solution then
raised In pit to 5. Eh
to >1000 nv with C12.
pH maintained >4 and
Eh near 1000
Starting Solution adjusted
to pll • 5. filtered
aqueous sample (most
of chromium In solids)
One Hour Sample, exposed
only periodically to
Cl?, Eh maintained at
1000 mv and pll >4,
filtered aqueous sample
(time to 2 hrs. showed
no Improvement in Cr
oxidation)
Chromium Slurry Oxidation
3.5 liters of above
oxidized slurry reoxidized
by constant exposure to
Concentration (opl)
Cu
7«
Cr
Nl
Cd
Al
-------
TABLE R.I26. CONTINUED
Sample No.
2372
2373
2347
Condition
Thirty minute exposure
to C12, filtered sample
One hour exposure to
to Cl2. filtered
sample
One and one-half hour
exposure to CI?,
filtered sample, most
Concentration (gpl)
Cu _£SL to _Cr._ Ni Cd Al
O.OS <0.01 0.09 1.36 2.37 0.28 0.01
0.06 <0.01 0.11 1.S9 2.49 0.28 0.01
.
.
cr
0.07 0.01 0.11 1.6S 2.31 0.24 0.01 bT?4
dizc'd)
solids In solution,
final filtrate
Chromium Precipitation
Lead sulfate (2X stolchto-
oetric requirement) added
to solution 2374 to precipi-
tate PbCr04> pH maintained at 4. 10 liter*
2375 Thirty minute exposure 0.07 <0.l. 0.11 0.009 2.11 0.22 <0.01
2376 Final Filtrate 0.06 <0.01 0.10 0.006 2.03 0.21 <0.01
Chrajl un .Pi"eci£Ua£(on f 2_
IS liters of solution 2340 (only 42 I oxidized) exposed to lead su'fate without
reoxidizing.
-------
TABLE 8.126. CONTINUED
2348 Starting Aqueous Solution; 0.02 « D.L. 0.08 0.73 2.31 0.28 < D.L.
pH adjusted to 4.0
2349 One-half hr exposure 0.02 < 0.001 0.08 0.56 2.32 0.20 <0.001
2352 One hr exposure 0.02 < D.t. 0.08 0.17 2.31 0.28 < D.I.
2364 Lead Chrooate-Lead Sulfate < D.L. < D.L. < D.I. 7.51 0.008 0 04 < D L
Residue washed, dried, ' '
redlssolved In 10 v/o acid;
46.4 g leached In lOOcc solution.
Nickel Precipitation
Na2S added as a solution over a period of 20-30 ninutes. maintaining pH • 5; approxlnalely
stolchlonietrlc amount of Na2S added to 10 liters
2367 Starting Solut1on(2352 not 0.02 » 0.01 0.07 0.16 2.27 0 26 < D I
2376) ; Initial pH • 3.7 ' '
2378 Final Filtrate . « D.L. « D.L. « D.L. 0.04 0.07 < D.L. « D.I.
(Cl- • 2.61 gpl. SO^ • 28.23 gpl)
NOTE: 'Barrel 1 sludge composition (34.401 soll-Js. 65.61 IL^O): 7.8 Cu; 18.3 Fe; 11.5 Zn; 1.2 Cr; 5.5 HI; 2.8 At
•Standard leach on 35.58 pounds of sludge (55°C»; final volume 29.6 liters of 20 gpl Fe; diluted to
57.6 liters, pll - 1.9.
•Conditions changed to precipitate potasslua jarosite; temperature adjusted to 90°C; pH to 2.45;
1 gn K2S04/gm Fe. 6 hours.
•Jarosite solution set 8 hours (overnight) then solution decanted off; residue diluted then filtered
in filter press.
-------
TABLE 8.126 CONTINUED
•Jaroslte solution diluted to decrease Zn content to design range; total volume approximately 90 liters.
•Cu SX: LIX 622 (10 v/o); 2 stages of extraction (0/A • I); 2 stages of strip (150 gpl H?S04 (0/A •!).
flow rate 250 cc/otn., InlttalpH • 2.0. final rafftnate pH • 1.). temperature 30-40%. Contacts
perforated In I gal. mtxer-1 gal. settler systeo.
•Zn SX: DEHPA (40 v/o); 4 stages of extraction (0/A - 1); pH adjusted after first two contacts back
to pll • 2; 3 stages of strip (200 gpl ^$04, 0/A • 1); temperature 30-40°C; Initial pH • 2.0,
final raffinate pH • 1.3.
•Zn SX: OCHPA (40t) test repeated at Initial pH - 3; readjusted pH after second contact to
pH • 2. Aqueous solution froa previo-* test '90 liters) used. Otherwise, conditions same as
u above.
01
•Cr oxidation by Eh control tried on 30 liters of solution. Results required reoxidation, 10 liters.
of slurry reoxldized In flowing Clj (0.2 1/ain.) while pH ulntalned greater than 4.
•fbS04 used as precipitating agent for chronlun renoval as PbCrfy. PbCK>4 can be redlssolved to form
chromic acid and lead sulfate regenerated for reuse. Very effective precipitant and easily filtered.
•Na2S used as nickel precipitant. Very rapid precipitation but pH must be maintained near 5; If lower
H2$ odor results. If much higher then nickel precipitates as NI(OH)2.
-------
Elements Present: S1, Fe, K, Cr, Ni, Cu, Zn, S, P.
Figure 8.26. Sequential test series three crud: qualitative analysis.
436
-------
8.13.5. Sequential Test: Series Five (111
Purpose of lest: Large Scale leach to carry out all unit operations.
Results
Comments
Table 8.127. Sequential Test: Series Five
Table 8.123, Sequential Test: ton Recoverable Elements
Table 8.128, Reagent Consumotion for the Treatment of
50.6 kg (111 pounds) of Metal Finishing
Hydroxide Sludge
Table 8.129, Acid Leach of Residue - Jaroslte Solids:
Sequential Series Five
Table 8.130, Toxicity Test Applied to Releached Jaroslte
Product from Sequential Test Five
The entire sequence of operations went well:
The large leach system easily handles 100 pounds of
sludge material. The solid-liquid ratio can be varied
to produce a solution that contains between 10-15 gol
Fe. This iron level is required for production of an
easily filterable jaroslte-residue mixture.
The Jaroslte In-situ precipitation produces a solid that
settles rapidly. Therefore, most of the solution can be'
decanted from the so'.ids. The slurry remaining can be
conveniently handled in the LASTA filter press. Irun
content can be decreased in the mixed metal solution to
0.5-1.0 gpl. This level 1s appropriate and can be
removed during Zn SX. The exit pH from jarosite
treatment is at a convenient level for Cu SX.
The sequential five test showed high copper loss to the
resldue-jarosite solids because of two reasons; the pH
of the leach *as high so copper leached into solution
was lower than usuai and the pH of the jarosite
precipitation was also higher than normal: *2.9 wnich
meant that some copper was precipitated. A leach of the
resldue-jarosite solids showed recovery of 75% of the
copper; Table 8.129.
Copper solent extraction in the Reister System operates
well. Interfaces are controllable and very little crud
forms. That which does form can be siphoned oft.
Copper contents can be handled up to at least 10 gpl.
Fifteen to twenty-seven liters/hr. can be treated.
437
-------
Gl
00
TABLE 8.127. SEQUENTIAL TESTS: SERIES FIVE (111.6
Sample No. Condition
2492
2493
2494
2494
2496
2499
25008
2525
2526A
2524
2526B
2527
Standard Leach (Barrel 18)
ZIZ liters
Jarosite Preclpltat'on
One hr exposure. 211 lit.
Seven hr exposure. 200 lit.
Final pH • 1.91
Cu SX (15 v/oLIX 622)
Starting Solution. Initial
pH • 1.91. 160 liters.
Raffinate, 2 hrs
Final Composite Raffinate
Final pH « 1.3
Final Strip
In SX (40 v/o DEHPA)
Starting Solution, pll • 2.
Raff, after stage 2.
Initial pll - 2; 160 liters
Raff, after stage 4.
Initial pll - 2.5
Composite Raff. iFlnal
pH • 1.3; CT • 1.36 gpl;
Final Strip Acid
Cu
3.25
2.29
, 3.05
3.05
0.01
O3
23.17
0.04
0.04
,
0.02
0.04
SO. • 46
q0.03
Fe
9.73
2.42
DTT7
0.57
0.55
0.52
-------
:TA8LE 8.127. CONTINUED
Sample No. Condition Composition (gpl)
•
2S64
2574
2580
*>
1*1
!o 2589
2592
2638
Cu Fe
Chromium Slurry Oxidation
Starting Solution. 75 lit., 0.04 0.22
Eh • 350 mv. p!l -1.30
Two hr exposure to chlorine, 0.03 0.08
Eh • 888 DV. pll • 4.38
Two and one-half hr expos- 0.03 0.12
use; Eh • 1005 nv, pll • 4.2
Four hr exposure; Eh • 0.02 D.I.
1138 ov. pil • 4.0
Five hr exposure. Eh • 0.03 O.L.
1132 rav, ptl • 4.2 (CP- 12.6 gpl; SOa* • 33
Leach of solids fron 0.004 3.21
oxidation; 4.98 9 (n
lOOrc of 10 v/i- H2S04
Chroralua Precipitation
Zn Cr Ni
0.07 2.67 1.75
0.05 1.70 1.45
<52TWof the
chromium oxtd. )
0.06 2.02 .58
(66TWoxid.
0.05 2.2* .76
(B2:5Y~oxld.
0.06 2.28 .68
.9) (83V«~ox1d.
-------
*»
««
o
TA3LE 8.127.
CONTINUED
Sample No. Condition
2640
2641
2643
Cu
Lead Sul fate-lead Chromate Re leach
Leach of solids. 50g < D.L.
solids In lODcc 20 v/o H.SO.
Wash of solids from ' 4< D.L.
2640 (154 cc)
Hash of solids frea < D.L.
2640 (458 cc)
HIS Precipitation
Fe
0.15
0.07
0.002
Na?S added (2X stotchionetrtc addition as
45 Dinute period, pll maintained between 4-5
2605
2606
2607
2609
2610
2644
2645
2646
Starting Solution 2603 0.04
Ten rain, exposure < O.L.
Twenty min. exposure "
Forty-five rain, exposure. "
Na2S addition complete.
Sixty rain, exposure •
Ion Exchange of Final Solution
NiS filtrate, feed to < 0.001
column; IRA- 200. 150 cc
IX of 2644 (H fom)
IX of 2644 (Na fora)
-------
TABLE 8.127. CONTINUED
Leach
Sludge leached to produce an Iron content of approxlutely 10 gpl, standard conditions.
Jaroslte
Standard conditions: 88-92°C. pH - 2.2-2.7. 1 g K,SO./g Fe. 7 hours.
Solution slurry set overnight then solution decanted off. residue diluted then filtered
In filter press. Solids subjected to EP test.
Cu SX
l~nrS22 (IS v/o); 2 stages of extraction. 0/A - 1; 2 stages of strip. 150 qpl H.SO.;
0/A - 1; temperature 40 - 50°C; flowrate 250 cc/nln etch phu$e; Initial pll:- 1.9; *
contacts perforned In Relster system.
In SX
DTliPA (40 v/o); 4 stages of extraction. 0/A • 1; pH adjusted after first two contacts
to 2.5; 3 stages of strip. 200 gpl H?SO., 0/A • 1. temperature 30-40°C; Initial
pll • 2.0; final rafflnate pll - 1.3. z 4
Chropilua Slurry Oxidation
pll maintained between 4-5; chlorine sparged Into vessel at 5 liters/din; Eh >1000mv;
system agitated to suspend solids In solution. Degree of chromium oxidation deter-
mined by filtering sample, exposing solution to IRA 900 anionlc exchange resin to
remove oxidized chromium species; degree of oxidation calculated by difference.
Chrcroiurn Precipitation
Lead sulfate (=2X stoichioaetric requirement) added to agitated solution. PH° 4
maintained. Lead chromate solids filtered easily.
HiS Precipitation
Sodium suiflde solution (325 gpl) added slowly to solution over a period of 45 nln.;
Solution agitated to suspent particles; pll • 4-5. No odor problen.
Ion Exchange
Most of the final solution can be recycled as make-up Mater.
-------
TABLE 8.128. REAGENT CONSUMPTION REQUIRED FOR THE TREATMENT OF 50.6
KG (111.6 POUNDS) OF METAL FINISHING HYDROXIDE SLUDGE
Unit Operation
Reanent
Amount
Copper
•Zinc
Ch roml ugLpx 1 datl on
Chlorine
Electrochemical
Leach
Jaroslte Precipitation KOH
H2°2
Filter Press HjO
Solvent Extraction
LIX 622
KERMAC 4708
D2EHPA
KERMAC 470B
H2S04
na
NaOH
CU
Chromium Precipitation NaOH
PbS04
Nickel Precipitation Na2S
13.3 Kg (Concentrated acid)
12 liters new water
156 liters recycle water
10 liters (500 gpl)
2.5 liters (30 v/o)
14 liters
1.2 liters (one time addition)
6.8 liters • •
8 liters (Recycle add
150 gpi)
10.8 liters (one tine addition)
16.2 liters ( • •
11 liters (Recycle aclc,
200 gpl)
4 liters (4N)
8 liters (500 gpl)
Not established
Regenerates acid
1 liter (400 gpl)
4.2 Kg (one tine addition)
6 liters (325 gpl)
1 liter (ZOO gpl)
442
-------
TABLE 8
Sample
2698
2701
2702
Notes:
TABLE 8
Sample
2711
2712
2713
.129. ACID LEACH OF RESIDUE-JAROS1TE SOLIDS: SEQUENTIAL SERIES FIVE. VARIABLE
No. Condition
Initial pH • 0.5
Initial pH • 1.5
Initial pH • 2.5
. 10 grans solid slurried In
temperature. 18 hours.
. Solid starting composition
1.5 Al
Pll
Recovery Fran Solids(S)
Fe Cu Zn Cr
11.5 75.0 66.7 18.8
5.9 25.0 33.3 1.2
0.2 7.1 10.0 0.3
100 cc solution, pll adjusted to
(S): 19.8 Fe. 2.8 Cu. 0.28 Zn.
Hi Cd
100.0 < D.L.
0.0 < D.L.
0.0 < D.L.
Al
13.3
2.0
0.0
desired value; ambient
3.2 Cr. 0.04 HI. 0.0 Cd.
.130. TOXICITY TEST APPLIED TO RELEAQIEO JAROSITE PRODUCT FROM SEQUENTIAL TEST
No. Condition
Test One. pH • 3.24
Test Two. pH • 3.23
Test Three, pll • 3.34
FIVE
Concentration (mn/liter)
Fe Cu Zn Cr
5.73 4.23 1.94 0.55
5.10 4.17 1.99 0.54
4.19 1.89 9.01 0.46
Ni Pb
0.33 < D.L.
0.35
0.42
Al
1.68
1.68
1.27
Notes: . Test performed according to EPA designated EP Toxlclty test(27). EPA designated concent-
ration of contaminants for characteristic toxicity (mg/1): 1.0 Cd. 5.0 Cr, 5.0 Pb.
-------
Zinc solvent extraction In the Reister system using pH
adjustment after two contacts works well. Interfaces
are stable and controllable at 250 cc/min. (prooably
also to 450 cc/min.) Crud is not a problem. Gypsum
forms in the strip cell but can be filtered and not
returned to the extraction circuit (this is a bleed for
CaT* from the system). Iron is co-extracted with'zinc.
Zinc can be stripped by H2s.°d' iron is not stripped. A
bleed stream can be taken from the organic and treated
with KC1 to strip the iron. The resulting D-EHPA can
then be recycled to the zinc extraction circuit.
Aluminum is co-extracted and provides a means to
partially remove it from the system. A part of the
aluminum is stripped into the H.SO., a part into the
HC1. z *
Chromium oxidation is a slow process. Better Cl_
contact would accelerate this process. Electrooxidation
may be an appropriate substitute. Slurry oxidation
produces a small amount of solids, primarily Cr(OH)..
This solid can be recycled to the Initial leach system.
Chromium can be effectively stripped from the solution
after oxidation by uss of recycled PbSO.. The solid
formed Is easily filtered or settled from solution. The
lead chromate can be releacned to produce a concentrated
chromic acid solution. The rate of precipitation is
rapid, therefore, a small reactor can be used.
"Nickel can be stripped from the final solution by use of
a Na_S solution. The final sulfide treatment also helps
to scrip residual cations from solution.
The solution after nickel removal can almost entirely be
recycled to the leach-jarosite steps as make-up water.
The sulfide precipitation is rapid. Therefore, a small
reactor can be used.
8.14. TEST ASSEMBLY EQUIPMENT
8.14.1. Unit Operation Equipment
A list of the equipment in the test assembly is presented in Table 8.131.
The equipment list is organized according to the unit operations specified in
flowsheet Figure 6.1.
444
-------
8.14.2.. Pictorial Presentation of Test Assembly Equipment
A series of photographs of the test assembly equipment Is presented
-------
3. Diaphragm Plate - The diaphragm plate is a steel frame
around two rubber diaphragms. When clamped against the
filter plate, it forms the second half of the filtering
chamber. The rubber surface is ribbed similarly to the
filter plate to allow the fiitrate to drain. When filled
with water under pressure, the rubber expands, compressing
the solids in the chamber.
4. Moveable Head - The moveable head is attached to the
cylinder rod. It distributes the force of the hydraulic
cylinder to create the clamping pressure on the plates.
5. Filter Cloth - The two panel filter cloth is a
polypropylene weave which is hung between the filter and
diaphragm plates. The slurry is pumped into the press,
between the two panels. The weave of the cloth retains tne
solids and allows the filtrate to pass through to drain.
6. Hydraulic Pump and Cylinder - A double acting hydraulic
cylinder is mounted in the rear fixed head. The cylinder
rod extends and retracts to open, close, and clamp the
plates. Oil to drive the cylinder is provided by a hand
pump on top of the rear fixed head. The pump is equipped
with a valve to direct the oil to the rod or head end of
the cylinder.
Solvent Extraction
'Two Keister ten-cell solvent extraction testracks. Each cell has a
one-gallon mixing chamber and a one-gallon settling chamber. Each
mixing chamber is agitated with a one-seventh horsepower variable
speed motor. Solution flow is conrolled by the agitator speed and
•its position over the solution inlet opening. Mixed solution
continuously overflows a weir into the settling chamber. Organic
phase separates to the top of the settling chamber and overflows a
weir to an organic chamber. The aqueous phase and organic phase
both flow continuously from the settling chamber.
'Associated solution pumps (flowrate adjusted to 500 cc/min.) to
supply the aqueous feed, loaded organic, and strip solution feed to
the SX chambers. Blue White C1760LP.
'Two hundred liter polyethylene storage vessel for collecting the
raffinate from testrack; two required, one for raffinate from Cu SX
and one for raffinate from zinc SX.
*ph controller (Cole-Parmer Model K-5660-00) for adjusting the
aqueous ohase ph from stage two of the four stage zinc SX set-up.
Copper Sulfate Crystallization
•Two liter reaction kettle with four-neck ground glass top and
closed stirrer system. System will treat one-fourth to one-half of
646
-------
strip solution exiting the copper SX system. Anticipated treatment
of ten percent bleed stream from strip solution; recycle of acid
from copper sulfate crystallization cell to strip circuit.
'Solution pump (flow adjustable to 250 cc/min.). Blue-White Model
C1760LP.
Copper Electrowinning
'Lambda LES-F regulated power supply, 0-9 v, 0-100 amp.
'Laboratory scale electrowinning cell, 13-Inch by 6-Inch by 8-inch
chamber. Twenty-one electrode slots for 6.75-inch by 5.25-inch
.electrodes. Copper cathodes, lead anodes.. System will treat strip
solution at a current density of 20 amp/ft at 2.5 volts. Copper
content decreased by 5-10 gpl.
'Solution pump Blue-White Model C1760LP.
Z1nc Sulfate Crystallization
'Two liter reaction kettle with four-neck ground glass top and
closed stlrrer system. System will treat up to one-sixth of strip
solution exiting zinc solvent extraction sy«tem. Anticipated
treatment of of ten percent bleed streams from strip solution;
recycle of acid from zinc sulfate cyrstallization eel1 to strip
circuit.
•Solution pump; Blue-White Model C1760LP (to 250 cc/min.).
Chromium Oxidation
Chlorine Oxidation
'Two hundred liter nalgene tank with vented top cover.
'One-fourth horsepower direct drive 115 v agitator with
three-fourths Inch diameter, 36-inch long epoxy coated shaft with
four-Inch impeller.
'Sparger for chlorine gas dispersion in solution slurry.
*pH monitor, Orion 601A.
'Solution pump for reagent addition, Blue-White Model C1760LP.
'Two 100 liter nalgene canks.
'Two one-fourth horsepower direct drive 115 v agitators with
three-fourths inch diameter, 36 inch long epoxy coated shaft with
four-inch impeller.
'Chlorinator assembly.
447
-------
*pH controller. Cole Partner Model
'Chlorine tank.
Electrochemical Oxidation (This cell is not of sufficient size
to treat a day's production of solution).
'Lambda LES-F regulated power supply. 0-9 v. 0-100 amp.
'Laboratory scale electrochemical cell. 13-inch by 6-inch by 8-inch.
Two cation selective membrane dividers to separate anode chamber
from cathode chamber. Lead anode, copper cathode. Volume and
number of electrode chambers variable.
'Solution pump for anolyte recycle. Blue-White C1760LP.
"Plexaglas anode chamber with nafion membrane sides. Larger
plexaglas cathode chamber.
'Circulation pumps, Masterflex double head, blue white catholyte
recycle pump (C1760LP).
'Lambda LES-F regulated power supply, 0-9 v. 0-100 amps.
'Electrodes, solid lead sheet and lead wool sandwiched between two
perforated lead sheets.
'Storage reservoirs, two. 30 liter nalgene tanks.
Chromium Precipitation
'One hundred liter polyethylene reactor vessel.
'One-half horsepower air driven direct drive agitator for variable
speed control, 316 S.S. shaft, 36-inch length, 3/4-inch diameter
shaft, 4.5-inch impeller.
*pH monitor, Orion 601A.
'Solution pumps for reagent addition and solution transfer.
Nickel Precipitation
'Two hundred liter nalgene tank.
'One-fourth horsepower, 110 v agitator. 316 S.S. One-inch diameter
shaft, 36-inch long, two 8-inch diameter impellers.
*pH monitor, Orion 601A.
'Solution pump to provide reagent addition, Blue-White C1760LP.
'Solution pump for recycle to leach solution as make-up water,
Cole-Partner.
468
-------
Figure 8.27. Leach-jarosite test system.
-------
•vMtotol* copy.
«
Figure 8.28. LASTA filter press.
450
-------
I
Rcp.odufrd from
copy.
Figure 8.29. Small scale continuous solvent extraction system.
(Bell Engineering 600 CC system)
451
-------
Figure 8.30. Reister one gallon mixer-settler continuous solvent
extraction system.
452
-------
Figure 8.30. Reister one gallon m1xer-s.:ttler continuous solvent
extraction system.
453
-------
Figure 8.31. Chromium oxidation by chlorine sparging.
454
-------
Figure 8.32. Chromium oxidation by chlorinator system.
455
-------
Figure 8.33. Electrochemical oxidation of cnromiura.
456
-------
Figure 8.34. Lead chromat-i precipitation.
457
-------
8.15. DETAILS OF ECONOMIC ANALYSIS
An economic analysis was presented in Section 6.4 for a 50 ton per day
facility. The mass flows, equipment lists, factored capital cost data sheets,
and operation cost estimates are presented in this section.
8.15.1. Leach-Jarosite Precipitation Filter
The equipment list is presented in Table 8.133 and the factored capital
cost (FCC) is presented in Table 8.134. The operations cost is presented in
Table 8.132. The total operation annual cost for this series of unit
operations is $343,000; 8.4 cents/lb. of jarosite plus leach residue.
8.15.2. Jarosite Ponding
The jarosite ponding cost is estimated from recent cost data on a state of
the art mineral prqcessing tailings pond. The result is presented in Table
8.125; the factored capital cost is $390,500, the FCAC is $108,200 and the
operations cost is $25,400. The total annual cost is $133,600; the cost per
pound is included with the cost for leaching and jarosite precipitation, i.e..
8.4 cents/lb.
8.15.3. Copper Solvent Extraction Electrowinning
Copper solvent extraction costs are estimated from data presented by
Wood' . The equipment (FCC and FCAC are presented in Table 8.136. The
operation cost was presented previously in Table 8.131. The total annual cost
for this series of unit operations is $299,000; 80.2 cents/lb. of copper
produced. This cost is approximately the same cost, as the operating cost
incurred by a current copper smelter (operating cost only, not including any
capital cost). The estimated cost is greater than the current value of the
copper product.
8.15.4. Zinc-Iron Solvent Extraction. Zinc Sulfste Crystallization
Zinc and iron solvent extraction are estimated from data presented by
ctore
458
Wood* . The equipment cost and factored capital cost are presented in Table
-------
TABLE 8.132. OPERATING COST SUMMARY
Leach-Jarosite PrecipUafion-FiIter
COST(S/Yr)
1. Reagents
Acid: 1030 gal/d / 11.000
KOH: 446 gal/d 47,200
Steam: available
2. Labor
5375/week (Oct. 9, 1984 Mall Street Journal average
weekly pay) plus 30% benefits: S25350/man
2 persons/shift; 3 shifts 152,100
3. Maintenance
68 of Factored Capital Annual Cost (FCAC) 7,200
4. Power
5X of FCAC 6,000
TOTAL 223,500
Jarosite Storage
1. Labor
1 person, 1 shift • 25,350
Copper Solvent Extraction
1. Reagents
Lix 622: 150 gal $6.840
KERMAC 470B 920
Acid, 330gal @ ISOgpl H2S04 15
Total 7.775 (One time cost, included
under capital cost}
Organic loss: 13 mg/1 9,200
2. Labor
2 persons/shift. 3 shifts 152.100
459"
-------
TABLE 8.132 CONTINUED
Chromium Oxidation. Precipitation, and Recovery
1. Reagents
Acid: 60ppd H2S04 600
NaOH: 0.7 tpd 69.300
PbSO.: 2.52 tons (Start up only) $4,300 (Included
In capital)
2. Labor
3 persons/shift; 3 shifts 228,200
3. Maintenance
6% FCAC 30,200
4. Power ;
5X FCAC 54,200
Electrooxidation,' 3755 kwhr/tonne 25,200
TOTAL 407,700
Nickel Sulfide Recovery
\
1. Reagents
Caustic: 29ppd 1,400
Phosphate: 7300 pounds (One time cost, 3200)
Na2S: 0.51 tpd 69,000
. H2S04: 35ppd 400
2. Labor
2 persons/shift; 3 shifts 152,10U
3.Maintenance
5X FCAC 3,900
4. Power
5% FCAC 3,200
TOTAL: 230,000
460
-------
TABLE 8.132 CONTINUED
COST (S/Yr)
3. Maintenance
61 FCAC 5,200
4. Power
5SFCAC 4,300
Electrowinning, 2500 kwhr/tonne Cu 35,100
TOTAL 205.900
Zinc and Iron Solvent Extraction
1. Reagents
OEHPA: 790 gal $16,850
KERMAC 510: 1190 gal 1,300
HC1: 700 pounds 450
HgS04: 1290 pounds 100
Anterlite LA-2: 320 gal 8,000
267700 (One time cost.
included in capital
cost)
Organic loss: 13 mg/liter 4,100
Caustic: 0.057 t/d 5,600
2. Labor
3 persons/shift; 3 shifts 228,200
3. Maintenance
61 of FCAC 9,700
4- Power « i««
5X FCAC 8.100
25* FCAC for crystallizer power 14,000
TOTAL 269,700
461
-------
TABLE 8.133. LEACH-JAROSITE PRECIPITATION-FILTER EQUIPMENT LIST
Feed System
1. Vibratory feeder; C » 965(20 ft2)0'559:
2 tph, 1 each. n ,ft
2. Recycle solutio;. feeder; C » 156(30gpm)u
-------
TABLE 8.134. FACTORED CAPITAL COST FOR LEACH-JAROSITE PRECIPITATION-
FILTER SYSTEM.
Cost (S. MSS * 794)
1. Purchased Equipment Costs 141,500
2. Installed Equipment Costs (1.40 X Item ') 198,100
3. Process Piping (302 of 2) 59,400
4. Instrumentation (102 of 2) 19,800
5. Auxiliaries (52 of 2) 9.900
6. Outside Lines (55 of 2} 9.900
7. Total Physical Plant Costs (Sum of 2 through 6) 297.100
8. Engineering and Construction (202 of 7) 59,400
9. Contingencies (152 of 7) 44,600
10. Size Factor (Small Coimorcial, 102 of 7) 29,700
11. TOTAL PLANT FIXED CAPITAL COSTS 430,800
(1) Format from Mineral Processing Equipment Cost
and Preliminary Capital Cost Estimations", E.A.
Parkinson and A. L. Mular, Canadian Institute
Mining and Metallurgy, V. 18. 1978.
YEARLY COST, Based on 60 Month Pay-Off $ 119.500
Period, 122 Interest
YEARLY OPERATING COST $ 223,500
TOTAL YEARLY COST $ 343.000
463
-------
TABLE 8.135. JAROSITE PONDING EQUIPMENT LIST
The cost for jarosite ponding storage is estimated by determining
the necessary pond capacity. Assumptions include: la.nd is available,
pond capacity great enough for ten year storage, controllable access
only. The cost is estimated by ratioing the capacity of a known
recent tailings pond as described by Jones (52 ); i.e., 36x10°(1980
cost) for a capacity cf 7.3xlOb cubic meters; includes pumping, pH
control, instrumentation, and monitoring.
The storage capacity needed for present estimate is 50,900 M^.
Cost « 6xl06 f 50,900/7.3x106 ] ' [ 794/620]
= S390.500.
«
This cost is estimated to be the current factored capital
cost. The annualized cost is SI08,200.
464
-------
TA3LE 3.136. COPPER SOLVENT EXTRACTION-ELECTROWINNING EQUIPMENT LIST
AND FACTORED CAPITAL COST
The capital cost tor the solvent extraction system is estimated
from fabricated equipment cost for solvent extractors presented by
Woods(50 ):
Data: Mixer-settlers
Size Size Range Cost(5) Exponent H&S
1.5 llt/s 1.5-10 Iit/s 7.00C 0.4 600
(Includes: installed mixer-settler, including explosion
proof motors, drives, and within module piping,concrete.
steel, instruments, electrical, insulation and paint,
and necessary labor.)
Factors for Materials
2.00 for 316 stainless steel
1.4 for rubber lined
2.0 for tankage and crud removal system
The capital cost for the present system is:
Cost = 57,000 [lit/s,present/lit/s.Woods ]°''"f (MSS,now/M4S,then)x
(No. of cells)(2.0 stainless steelH2.0 for tankage)
Cost = 57,000[1.86/1.5]°'4 (794/600M2M2H5) = 5201,900
The capital cost for the electrowinning system is estimated from
fabricated equipment cost presented by Woods (50 ):
Data: Size Size Range Cost(S) Exponent H&S
1010g/y 2-60xl09g/y SxlO6 1.0 600
(Includes: cells, transformers, rectifiers, and electrical
distribution)
The capital cost for the present system is:
Cost =[5xl06][1.69xl08/lxl010 ] ' (794/600) =5134,200
The total factored capital cost is, therefore, 5336,100
The FCAC is ' 93,100
The yearly operating cost is 205,900
The total yearly annualized cost is 5299,000
465
-------
8.137. The total amount of operations cost for this series of unit operations
is 5453,000; 43.0 cents/lb. of zinc sulfate produced. This estimated cost is
greater than the current value of the zinc sulfate product.
8.IS.5. Chromium Oxidation. Precipitation and Chromic Acid Recovery
The equipment list is presented in Table 8.138 and the factored capital
cost is presented in Table 8.139. The total annual operations cost for this
series of unit operations is 5911.300; 119.6 cents/lb. A potentially lower
cost oxidation process is discussed in Section 6.4.
8.15.6. Nickel Recovery
The equipment list is presented in Table 8.140 and the factored capital
cost is presented in Table 8.141. The total annual operations cost for this
series of unit operations is $294,200; 49.9 cents/lb. Other alternate products
were considered and discussed in Sections 6.4 and 8.15.7.
8.15.7. Alternates
8.15.7.1. SQg-Og Chromium Oxidation, Nickel Oxide Production
The oxidation of chromium by.chlorine or by electrochemical mea.is Is the
most expensive unit operation in the sludge treatment flowsheet. If the cost
of this unit operation could be decreased then the overall ROI would be
Increased. The substitution of an SOg-Og oxidation system may prove to be a
much cheaper means of oxidizing chromium. This substitution has been discussed
in Section 6.4. The data for the substitution is presented in Tables 8.142,
8.143, and 8.144. The results on the overall ROI were presented previously in
Tables 6.J2a and 6.33a.
8.IS.7.2. Copper Cementation
Copper cementation as a substitute for copper solvent extraction-
el ectrowirning may be a more economical way to recover copper from a sludge
leach solution stream. Biswas and Davenport report that (based on Ranchers'
Exploration data) copper cementation by iron costs S0.35/kg copper less than
466
-------
TABLE 8.137. ZINC-IROM SOLVENT EXTRACTION EQUIPMENT LIST AND FACTORED
CAPITAL COST ESTIMATE
Cost data are not available for commercial zinc solvent extraction
facilities. The equipment required is, however, similar to that re-
quired for copper solvent extraction. The major difference is the
number of cells. Foi copper solvent extraction the orevious cost
was based on three stages of extraction, two stages of stripping. .
The zinc-iron solvent extraction is estimated based on the following
assumptions:
a. Solvent extraction of iron and zinc requires ten cells for
loading/stripping. The capital cost/cell (based on copper SX)
is $201,900/5 = 540,400. Therefore, for iron and zinc the
FCC = $404,000.
b. Zinc sulfate crystallization is estimated: 990 gal per day
of strip solution containing 140 gpl zinc. A batch crystal lizer
cost (57) 1s $19,000 ,700( factors
L J Table 6.29.)
c. The hydrochloric acid strip solution generated per day is
330 gal containing 18 gpl iron. A small pilot size SX plant to
recover the KC1 would be required using Amber lite LA-2 ( 34).
The volume of solution to be treated is very small (330 gpd)
compared to the volume of treated leach solution ( 42,000gpd).
Two mixer-settlers would be required for extraction and two
for stripping. The cost for four cells is estimated from Wood
to be:
Cost = 7000 [o.04 lit. /sec. /I. 5 lit. /sec.] ° [
794/600
= $9,300/cell. This includes mixer-settlers, explosion
proof motor, drive, piping, concrete, steel, instruments, electrical,
insulation, paint. Total cost for four cells = 537,200.
d. The concentrated ferric chloride may be a marketable product
but neither a credit not a penality is tak«n for disposal.
The FCC totil cost for solvent extraction, crystallization, and
stripping .of hydrochloric acid from the DEHPA strip solution is
$634,900. A one time cost for reagents is included in the capital cost,
i.e., 26,700. Therefore, the total capital cost is $661.600. The
The PC AC is 183.300
The operation cost (Table 8.1^«.s 269,700
The total annualized cost is $453,000
467
-------
TABLE 8.138. CHROMIUM OXIDATION. PRECIPITATION, AND RECOVERY SYSTEM.
EQUIPMENT LIST
COST(SPM&S=500)
1. Storage tank; C = 0.897(45,OOOgal)°'897: 26.800
Fiberglass, 2 each.
2. Electrochemical oxidation cells
3000 gal, 1500 amps, 15 units 480,000 Current
3. Precipitation vessels; C = 27.5(1040)°'629: 4,350
Stainless steel, 1040 gal, with agitator, 2 each.
4. Filter drum; C = 8235M9)0'292: 38,900
19 ft2 filtration area, 2 each.
5. Releach vessel
ss, 100 gal, with agitator, 1 each 1,000
TOTAL (excluding 2) 71,000
TOTAL (M&S =794) 112.800
TOTAL (including 2) 592,800
TOTAL (including one
time cost of
lead sulfate) 597.100
468
-------
TABLE 8.139. FACTORED CAPITAL COST FOR CHROMIUM OXIDATION.
PRECIPITATION AND RECOVERY
Cost (S. H&S ° 794)
1. Purchased Equipment Costs 597,100
•
2. Installed Equipment Costs (1.40 X Item 1) 835,900
3. Process Piping (30% of 2) 250,800
4. Instrumentation (105 of 2) 83.600
5. Auxiliaries (52 of 2) 41.800
6. Outside Lines (5S of 2) 41.800
7. Total Physical Plant Costs (Sum of 2 through 6) 1,253.900
8. Engineering and Construction (202 of 7) 250.800
9. Contingencies (15Z of 7) 188,100
10. Size Factor (Small Commercial, 10S of 7) 125.400
11. TOTAL- PLANT FIXED CAPITAL COSTS 1,818,200
(1) Format from Mineral Processing Equipment Cost
and Preliminary Capital Cost Estimations", E. A.
Parkinson and A. L. Mular, Canadian Institute
Mining and Metallurgy, V. 18, 1978.
YEARLY COST, Based on 60 Month Pay-Off $ 503,600
Period, 12% Interest
YEARLY OPERATING COST $ 407.700
TOTAL YEARLY COST $ 911,300
469
-------
TABLE fl.MO. NICKEL SULFIDE PRECIPITATION SYSTEM EQUIPMENT LIST
COST (S(3M&S=500)
1. Precipitation vessels; C = 27.5{1040)0'629: 4,400
Stainless steel, 1040 gal, with agitator, 2 each.
2. Filter drum; C = 8235(19)°'292: 38,«*00
19 ft2 filtration area, 2 each
3. Releach vessel;
ss, 100 gal, with agitator, 1 each 500
4. Precipitation vessel; C = 30.0(100)°'58: 500
Rubber lined stainless steel, 130 gal, with
agitator, pressure vessel. 1 each
TOTAL 45,900
TOTAL (M&S = 794) 72,900
TOTAL (including one
time cost for
phosphate) 76,100
470
-------
TABLE 8.141. FACTORED CAPITAL COST FOR NICKEL RECOVERY
Cost (S. M&S = 794)
1. Purchased Equipment Costs 76,100
2. Installed Equipment Costs (1.40 X Item 1) 106.500
3. Process Piping (302 of 2) 32.000
4. Instrumentation (102 of 2) 10,600
5. Auxiliaries (52 of 2) 5.300
6. Outside Lines (52 of 2) 5.300
7. Total Physical Plant Costs (Sum of 2 through 6) 159,700
8. Engineering and Construction (202 of 7) 31,900 .
9. Contingencies (152 of 7) 24,000
10. Size Factor (Small Commercial, 102 of 7) 16.000
11. TOTAL PLANT FIXED CAPITAL COSTS 231,600
(Sum of 7 through 10;
(1) Format from "Mineral Processing Equipment Cost
and Preliminary Capital Cost Estimations", E. A.
Parkinson and A. L. Mular, Canadian Institute Mining
and Metallurgy, V. 18, 1978.
YEARLY COST. Based on 60 Month Pay-Off $ 64.200
Period, 122 Interest
YEARLY OPERATING COST $ 230.000
TOTAL YEARLY COST $ 294.200
471
-------
TAHLE 1.142. INCO S02'02 OXIDATION EQUIPMENT LIST
COST(SgMSS=500)
1. Oxidation cells; C = 472(400)°-E2: 50,700
Flotation agitation cells, covered and vented,
400 ft?, 3 eacn
2. Drum Filter; C = 8235(19)0'292: 19,500
19 ft2 filtration area, 1 each
3. Rotary kiln
Includes refractory lining capable of 120QOC, firing
system, hot cyclone, water cooling, 6' diam., 8'
length, 30 tpd capacity (only need a 1 tpd capacity),
SI40,000 0 MSS 545, Fully instrumented. 128,400
4. Covered high temperature discharge conveyor
18"x25', S8.000 @ M&S 300 13.300
?. Bag collector
Small gas flows, cost includes motor and drive,
S2.500 3 M&S 300 4.000
TOTAL 215,900
TOTAL (M&S = 794) 342,800
A 72
-------
TABLE 8.143. FACTORED CAPITAL COST FOR S02-02 CHROMIUM OXIDATION,
NiO PRODUCTION SYSTLM
Cost (S. M&S = 794
1. Purchased Equipment Costs 342.800
2. Installed Equipment Costs (1.40 X Item 1) 479,900
3. Process Piping (302 of 2) 144.000
4. Instrumentation (10% of 2) 48,000
5. Auxiliaries (52 of 2) 24,000
6. Outside Lines (5% of 2) 24.000
7. Total Physical Plant Costs (Sum of 2 through 6) 719,900
8. Engineering and Construction (202 of 7) 144,000
9. Contingencies (152 of 7) • 108,000
10. Size Facto- (Small Commercial, 102 of 7) 72.000
11. TOTAL PLANT FIXED CAPITAL COSTS 1.043,900
(Sum of 7 through 10)
(1) Format from "Mineral Processing Equipment Cost
and Preliminary Capital Cost Estimations", E. A.
Parkinson and A. L. Mular, Canadian Institute Mining
and Metallurgy, V. 18. 1978.
YEARLY COST. Based on 60 Month Pay-Off $ 289,200
Period, 122 Interest
YEARLY OPERATING COST $ 484.600
TOTAL YEARLY COST $ 773.800
473
-------
TABLE 8.144. SOj - GZ OPERATING COST SUMMARY
COST($/Yr)
1. Reagents
Caustic: 47.0 tpd 14,100
Liquid S02: 103.400
2. Labor
2 persons/shift; 3 shifts 152,100
3. Maintenance
5X FCAC 17.400
4. Energy
51 FCAC , 14.500
Fuel for kiln, $6/1000 ftj 31,800
Electrical for kiln motors, 160 HP. 2 each,
0.08 S/kwh 151,300
TOTAL 484,600
474
-------
solvent extractlon-electrowinning. This comparison was made In 1975. If the
assumption is made that the cost difference then is the same as the cost
difference now, then the cost for producing 1130 pounds per day of cement
copper would be S180/day less than for producing electrowon copper by SX-EU.
This cost includes capital cost and operating cost. Therefore, the total
annual cost, presented in Table 6.30, of $299,000 should be reduced by $59,400;
the new total amount would be $239.600. However, the value of the cement
copper would be approximately one-half the value of electrowon copper. The
value of the electrowon copper was estimated to be $223,700. Therefore, the
value of cement copper would be $111,900. The difference between what is saved
in annual cost and the loss of value for the new product would, in fact, make
the ROI less rather than greater.
i
8.15.7.3. Production of CuSO.
An alternative to be explored is the relative cost of recovering
crystallized copper sulfate monohydrate instead of copper metal. If sne
assumes that the electrowinning unit operation has approximately tre same
annualized cost as the crystallization unit operation then the difference in
return would be the difference in the value of 1130 pounds per day copper
($0.60/lb.) and 3480 pounds per day of copper sulfate monohydrate (SO.75/1?.).
The yearly value difference 1s $637,600. This would change the ROI from 41 to
51 percent compared to the electrochemical oxidation flowsheet; and would
change the ROI from 69 to 83 percent compared to the SO^-O. oxidation
flowsheet. The cost differences are interesting enough so that the alternate
should be further considered.
8.15.7.4. Solvent Extraction of Nickel. Electrowinning of Nickel.
Precipitation of Chromiun Hydroxide, Production of Chromium Oxide
An attractive alternate is a major modification to the original flowsheet
that depends on the ability to selectively extract nickel from the leach
without removal of chromium. This appears to be possible by use of the
D.EHPA-EH& or LIX63-D.EHPA solvent extraction reagents. Further research is
needed to verify the conditions needed for such a system. Also, one should be
aware that'such a unit operation is more risky than previously suggested
475
-------
alternatives because SX of nickel by these reagents is not presently
commercially produced.
The alternate .flowsheet was presented previously in Figure 6.7b. The
equipment list and factored capital cost estimates are presented in Table
8.145; operating cost is presented in Table 3.146.
8.IS.8. Computer Mass Balances for 50 TPD Economic Analysis
The computer mass balance data for 'he 50 TPO economic analysis is
presented in Table 8.147. The computer mass balance program and software are
presented in a separate manual.
476
-------
TABLE 8.145. SOLVENT EXTRACTION OF NICKEL, ELECTROWINNING Of NICKEL,
PRECIPITATION OF CHROMIUM HYDROXIDE, AND PRODUCTION OF
CHROMIUM OXIDE
Solvent Extraction and Electrowinni.ig of Niekei
The FCC determined previously per mixer-settler cell for the flow
capacity of the present system was 540,400. The present system requires
five cells; 5201,900.
The electrowinning system FCC Is:
Cost = 6xl06 [l.72x!08gpy Ni/lxlO10 ] (794/600)
= SI 36.600
The FCC for SX and EU is: S338.500 (M&S 794)
The FCAC is: 93.800
Precipitation of Chromium Hydroxide and Production of Chromium Oxide
Equipment List CCST(S@K&S SCO)
1. Precipitation vessels: C = 27.5(1040)°'629: 4.400
stainless steel, 1040 gal, with agitator, 2 each
2. Filter drum; C = 8235(19)0t292: 19,500
19 ft2 area, 1 each
3. Rotary loin;
Includes refractory lining capable of 1?WC, firing
system, hot cyclone, water cooling, 6' diam., 8' length,
30 tpd capacity (only need Itpd capacity). S"40,000 @
M&S 545. fully instrumented 128.400
4. Covered High temperature discharge conveyor;
18"x25'. $8,000 P M&S 300 13,300
5. Bag Collector;
Small gas flow, cost includes motor ana drive,
S2.500 0 M&S 300 4,000
Total Equipment Cost (MiS
500) 169.600
Total (M&S 794) 269,300
FCC 819.800
FCAC 227.100
TOTAL FCC for SX, EH. Cr^C-j Production: 1,158,300
TOTAL FCAC for SX, EW, Cr^ Production: 320,800
477
-------
TABLE 8.146 OPERATING COST FOR TABLE 8.145 SEQUENCE
1. Reagents
LIX 63 (12.5%): 125 gal
DEHPA (16.0%): ISO gal
KERMAC 510: 715 gal
Organic loss: 13 mg/liter
Caustic: 74 tpy
2. Labor
2 persons/shift; 3 shifts
3. Maintenance
5* FCAC
4. Energy
5% FCAC ,
Fuel for kiln, $6/1000ftJ
Electrical for kiln motors, 160 HP, 2 each,
0.085/kwh
Electrowinning; 3755 kwh/tonne Ni, 173 tpy
COST(S/Y)
$6,200
3,400
800
10,400 (one time cost, included
in capital cost)
6,900
22.200
152,100
19.200
16,000
31.800
151.300
52.000
TOTAL OPERATION COST
$451.500
478
-------
TABLE 8.147. COMPUTER MASS BALANCE DATA FCR SO "?0 SLUDGE TREATMENT FLOWSHEET
se(**ss»tss«Bss*s*>sc*ttsstt««sst«t«st
COMPOSITE SLUDGE PRC3RAH
S*t«t*t««t*S**S****S*******SS«*SSSS3*t*
HERE IS THE COMPOSITION OF THE COMPOSITE SLUDGE (U/0>
CU NI CD ZN CR CA NA FE AL PB SI P
S.OO S.OO O.Oii S.OO S.OO l.OO l.OO 7.30 2.OO O.OO 2O.OO 2.OO
THE X SOLIDS IN COMP. SLUDGE t 23.OO
fH£ COMPOSITION OF THE COMPOSITE SLUDGE
GIVEN AS W/0 METAL HYDROXIDES
CA AND PB ARE GIVEN AS SULFATES
SI AND P ARE GIVEN AS OXIDES
CU NI CD ZN CR CA NA FE AL PB SI P
7.68 7.9O O.OO 7.6O 8.27 3.4O 1.74 12.O7 5.78 O.OO 42.78 6.17
THE TOTAL U/O OF THE HYDROXIDES, OXIDES AND SULFATES
•«tS IO5.34 *»«••
s*ts»s«s****«*sss«*s«s*s**«t**a>t*<*t
LEACH MODULE
ENTRY
(*s*8ts*
-------
TABLE 8.147. CONTINUED
EXTRACTIONS - X - ......
CU NI CD ZN CR CA NA FE AL PB SI P
9~. 70 9S.9O 100.00 93.10 96.50 15.00 1OO.OO 92. OO 96.90 O.OO O.OO 1OO.OO
X SOLIDS IN SLUDGE. DENSITY (G/CH3) OF SLUDGE
23. OO 3. SO
THE AMOUNT OF SLUDGE TREATED: (PPD) ... 1OOOOO.
T. EXCESS CAPACITY IN THE VESSEL, RESIDENCE TIME IN THE VESSEL (MRS.) ...
20.00 O.50
RESULTS ......... (INCLUDING RECYCLE SOLIDS/SOLNS)
THE ACTUAL AMOUNT OF SOLIDS TREATED i (PPD) 250OO.O
THL ACTUAL AMOUNT OF METALS IN THE SOLID: (PPD)
CU NI CD 2N CR CA NA FE AL PB SI P
J25O.UO 1250.00 O.OO 125O.OO 125O.OO 2TO.OO 23O.OO 1875. OO SOO.OO 0.00 5OOO.O
THE ACTUAL AMOUNT OF METALS EXTRACTED:
-------
TABLE 8.147. CONTINUED
AFTER THE LEACH
THE AMOUNT OF RESIDUE! (PPD) U236.7
THE ACTUAL AMOUNT QF METALS IN THE RESIDUE.- (PPD)
CU NI CO ZN CR CA MA FE AL PB SI P
78.75 SI. 23 0.00 61.23 43.75 212. SO O.OO ISO. OO IS. SO O.OO 5OOO.OO O.OO
THE W/O METALS IN THIS RESIDUE ....
CU N! CD ZN CR CA NA FE AL PB SI P
0.70 0.46 O.OO O.33 0.39 1.89 O.OO 1.33 0.14 0.00 44.30 O.OO
THE Z SOLIDS COMING FROM THE LEACH VESSELi 2.69
THE SOLUTION COMPt (GPL I
CU Nt CD IN CR CA NA F= AL PB SI P
;.4S 3.33 O.OO 3.SO 3.35 0. II 0.74 3.08 1.43 O.OO O.OO 1.47
THE TOTAL CPL IN THE LEACH SOLNi 72.37
THE AMOUNT OF EXCESS ACID IN THE LEACH SOLUTIONi (CPL) O.OO
I*********************** <«**(!**««* >*****•** «»*«t* •**(**(****?<•
SOLID/LIQUID SEPARATION
M ***«**»*****«**»**«** t*t*******C*S*********«******«****«**«*t*
STARTING CONDITIONS ..:..
AMOUNT OF SOLIDS ENTERING S/L SEPARATION: (<>Pt» 11236.7
AMOUNT OF LIQUID ENTERING S/L SEPARATION* (LI'l/DAY) 133927.
Z SOLIDS IN FIRST FILTER CAKE (NOT FROM REPULPS) I 70.00
* REPULP -- I
ZSOLIDS REPULP X GPL ACID
70. OO 5.00 60. OO
THE RESULT? ......
IXE AMOUNT OF SOLUTION EXITING FIRST S/L SEPARATION (LIT/DAY) 131945.
ITS COMPOSITION (GPL) .....
CU NI CD ZH CR CA NA FE «L PB SI P
3.43 3.33 O.OO -.SO 3.SS 0.11 O.74 3.08 1.43 O.OO O.OO 1.47
THE AMOUNT OF SOLUTION TRAPPED IN THE FILTER CAX.Ei (LIT/DAY) 1983.71
THE POTENTIAL LOSS OF METAL VALUES IN THE TRAPPED SOLN. »PPD)
CU NI CD :N CR CA NA FE AL PB SI P
13.O94 13.449 O.OOO IS. 320 13.345 0.483 3.222 22.231 6.244 O.OOO O.OOO 6.444
481
-------
TABLE 8.147. CONTINUED
limit '///////////////////////
IhE AHCUNT OF SOLUTION EXITING REPULP 1 IS I (LIT/DAY) 9824.1O
ITS COMPOSITION (GPL) ....
CU NI CD ZN CK CA NA FE AL PB SI P
0.38 O.59 0.00 O.58 O.59 O.O2 0.12 0.8S O. 24 O.OO O.OO 0.25
THE ?P& OF EACH METAL IN THE EXITING SOLUTION
CU NI CO 2N CR CA NA FE AL PB SI P
12.45" 12.751 O.OOO 12.645 12.831 0.399 2.639 18.549 S.154 O.OOO O.OOO 5.319
THE AMOUNT OF SOLUTION TRAPPED IN THE FILTER CAKEl (LIT/DAYl 2078.18
THE POTENTIAL I OSS OF KETALS IN THE TRAPPED SOLNs
CU NI CD ZN CR CA NA FE AL PB SI P
2.636 2.697 O.OOO 2.675 2.714 0.034 0.563 3.882 I.O90 O.OOO O.OvO 1.125
//// ////////.'///11/11/////////
THE AMOUNT OF SOLUTION EXITING SLSEP AFTER ALL REPULPS (LIT/DAY) 161767.
ITS COMPOSITION (GPL) ....
CU NI CD ZN CR CA NA FE AL PB SI P
3.28 3.35 0.00 3.33 3.37 O.IO 0.7O 4.83 1.36 O.OO O.OO 1.40
THE PPD METAL LOST IN THE FINAL FILTER CAKE I
CU NI CD ZN CR CA NA FE AL PB SI P
2.636 2.697 O.OOO 2.675 2.714 0.084 0.563 3.882. 1.O90 O.OOO 0.000 1.125
THE COMPOSITION OF THE FINAL FILTER CAKE
PPD OF SOLID ... 11236.7
PPD OF SOLUTION .... 4815.74
LIT/DAY OF TRAPPED SOLNi 2078.18
THE AMOUNT OF ACID EXITING IN THE FINAL SOLUTION
(IN GPL) 3.68
THE PPD ;•=• METALS EXITING FROM REPULPSI
CU ''I CD ZN CR CA NA FE AL PB SI P
12.46 f*.75 O.OO 12.64 12.83 0.4O 2.66 18.35 S.IS O.OO O.OO 5.32
482
-------
TABLE S.147. CONTIKUEO
GENERAL EXTRACTION PROGRAM : JAROS1TE PRECIPITAION
MMM«l**tMt«*t*MMMIt*S*i*M*»»S
OPL IN STARTING SOLUTION:
CU Nl CD 2N CR CA NA FE AL PB SI P
3.28 3.35 O.OO 2.33 3.37 w. 10 n. 7O 4.83 1.36 0.0-.. O.OO 1.4O
THE LlltSS/DAV STARTING SOLUTIONi 161767.
THE PERCLNT EXTRACTIONS]
CU NI CD /.N CR CA NA FE AL PB SI P
2.7 3.6 b.2 -T.O 15.O O.O O.O 97.O 44.5 10O.O 100.O 50.O
POUNDS IHHALd ENTER I Nli PROCESS:
CU NI CD /N CR CA NA FE AL PB SI P
lloB.o 1196.1 O.O 1186.1 1203.5 37.4 249.4 1721.1 4B3.4 O.O O.O 498.9
NEW GPL VALUEs - SOLUTION EXITING PROCESS f
CU Nl CD ZN CR CA NA FE AL PB SI P
3.19 J.23 v.OO 3.-Je» 2.67 w. 10 0.70 0.14 0.75 O.OO O.OO 0.70
THE MOLC. WT. OF THE PRECIP. SPECIESi
CU NI CD ZN CR CA NA FE AL *B SI P
97.5 92.7 146.4 99.4 1O3.O 110.O 0.0 5O1.0 7B.O 24B.O 64.O 65.0
THE MOLES OF PRECIP. SPECIES PRODUCED PER HOLE METALi
(XI NI CD ZN CR CA NA FE AL PB SI P
1.0 l.v> 1.0 1.0 1.0 1.0 l.O O.3 l.O 1.0 l.O l.O
THE PPD OF EACH PRECIP. SPECIES:
CU NI CD ZN CR CA NA FE AL PB SI P
48.4 6B.O O.O 36.1 357.5 «>.O 0.0 4492.B 621',9 O.O O.O 523.4
THE TOTAL PPD OF SOLID SPECIES PRECIPi 6148.1
GPL ACID IN SOLNi O.O
»»ttt»t»»«»»••*«»•«»»«»•«•»»»«»«»*»»«*«»»•»«*•*»»»»»»*«»*»*»•»*•
SOLID/LIQUID SEPARATION
tst*»is*tMsa>««««0«t^i*t3t«s«s«*>*«tMsi*t***$**«*t**»»*«**ee»»
STARTING CONDITIONS'
/
-------
TABLE 8.147. CONTINUED
AMOUNT OF SOLIDS ENTERiNS B/L S£PARATIONi (PPD) 6146. O6
AMOUNT OF LIQUID ENTERING S/L SCPARATIONi (LIT/DAY) 161767.
X SOLIDS IN FIRST FILTER CAKE (NOT FROM REPULPb)! 7O.OO
• REPULP -- 1
XSOLIDS REPULP X GPL ACID
70. OO l.OO 60.00
THE RESULTS ......
THE AMOUNT OF SOLUTION EXITING FIRST S/L SEPARATION (LIT/DAY) 16O6B2.
ITS COMPOSITION
-------
TA3LC 8.147. CONTINUED
THE AMOUNT-OF SOLUTION EXITING BLSEP AFTER ALL REPULPS (LIT/DAY) 161716.
ITS COMPOSITION (GPL) ....
CU NI CD ZN CR CA NA FE AL PB SI P
3.18 3.22 O.OO 3.25 2.86 0.10 0.70 0.14 O.75 O.OO O.OO O.7O
THE PPD METAL LOST IN THE FINAL FILTER CAKE I
CU NI CD ZN CR CA NA FE AL PB SI P
3.996 4.052 O.OOO 4.O85 3.595 0.13; O.B77 O.181 O.943 O.OOO O.OOO O.B77
THE COMPOSITION OF THE FINAL FILTER CAKE
PPD OF SOLID ... 6148.06
PPD OF SOLUTION .... 2634.SB
LIT/DAY OF TRAPPED SOLNs 1137.03
THE AMOUNT OF ACID EXITING IN THE FINAL SOLUTION
UN GPL) O.4O
THE PPD OF METALS EXITING FROM REPULPS:
CU NI CD ZN CR CA NA FE AL PB SI P
-.63 3.68 0.00 3.71 3.27 O.12 0.80.0.16 O.86 O.OO O.OO O.BO
****t*MtaMM*«*«9tlM«t**»MSt««$MM!)«*M*MM**9««***ttM«»
SX COPPER
t»**»**M**«M»****«*M«**»**«*»»*****»*»**M**»«*t«««**t**t*t*
THE SOLVENT EXTRACTION CONDITIONS ARE AS FOLLOWSI
TEMPERATURE (IN DEGREES CENTIGRADE)I 50
CONTACT TIME (IN MINUTES>1 3.O
O/A RATIO: l.OO ;
THE VOLUME FLOW KATE(L/D> OF ORGANIC SOLUTION RtOUIRED IS: 161716.
THE VOLUME X LIX 622 IN THE ORGANIC SOLUTION ISi 15.00
THE VOLUME X KERMAC 510 IN THE ORGANIC SOLUTION IS: 83.OO
THE VOLUME 7. 0 IN THE ORGANIC SOLUTION ISl O.OO
198.O SETS OF MIXING TANKS ARE REQUIRED, 3 STASES PER SET.
«*••»••»*****•*«««(«*»«*(*****«>*•***«»*«(****••(•««v»8*»**«a*s
THE PH FOR STAGE 1 IS 1.75
THE EXTRACTION EFFICIENCYS FOR STA5E 1 ARE:
CU NI CD ZN CR CA NA FE AL PB SI P
92.1O O.OO 0.00 O.OO O.OO O.OO O.OO l.OO O.OO O.OO O.OO 28.00
»SM*»**»M»»»»*MM*(*»««*****t*t**»«l****M**8t*l*lt»*tMM*«
THE AQUEOUS EXITING STAGE 1 CONTAINS (G/L):
CU NI CD ZN CR CA HA FE AL PB SI P
O.25 3.22 0.00 3.25 2.86 0.10 0.70 0.14 O.75 O.OO O.OO O.5O
kt*tf**»**«»****s«*a8**«t«e*i«*»»*«*«ts**t*«**s««««Ma*«s**(*«*
485
-------
TABLE 8.147. CONTINUED
M«**«e»«**«««8«tt****tMtl*Mt8«««MM*«tM»*t«M*»tt«Mt$*t»*
IHL* PH ru-r. STAGfc 2 15 1.90
!»K LXTRAiriON tFFliMENCYS FOR STAGE 2 ARE:
CD Nt CD !H CR CA NA FE AL f» SI f
Or-. .V O.J.I O.OO O.OO O.OO O.OO O.OO O.OO O.OO O.OO O.OO v.v'-O
I HI rUJUEOl:; (KITING ST.H3E 2 CONTAINS tb/LM
v*U M CO iN CK CA NA FE AL PB SI P
O..M 3. -•- O.OO 5.25 2. Bo 0. 1O O.7O 0.14 0. 73 O.OO O.OO O.SO
»*•«• «•••«»! «»»t »tf •«!(•••<• I («•«•»•••« «*««•*«»«« ««»««•(*«• «tt«
iXl M CO itt CK CA NA IE AL >'B SI P
...'.< .*..V ^..H> .i.OO O.OO O.OO -'.OO O.OO O.OO O.OO O.OO ->.OO
•«••«••••••«••«•••»•••••*•••••••»««•<•*•«•••••»•••••€•
«••*•*»•*«*••$•«<•»•*< * •t*»«e***«tt**«*»t»««*»<*»«tt«***«««a«*t
U<£ PH F'.'R SfACE .'IS 1.30
rHt fXTRACHON EFFICIENCYS FOR STAGE 3 ARE*
\'J M CO ZN CR CA NA FE AL PB SI P
45. JO v.OO O.OO O.OO O.OO O.OO O.OO O.OO 0.00 O.OO O.OO O.OO
•«•»«««« M«at«****»«<«t««****«t«S****«f »»$««»»»>«•••«» t»**»«»»t
THE AQUEPvS <.KITING STAGE 3 CONTAINS tG/L>l
•U SI CO ZN CR CA NA FE AL PB SI F
O..T ^.rr O.OO 3.25 2.8a O. IO -.'.7O 0.14 O.-'S O.OO O.OO O.SO
«««<«««t*«*«««««M*<*»*«**«*tt«*t»*tf «««»i «t««*f «•<•$»** tt«(»t*
vh£ ORGAN: r EXITING ST(>GE 3 CONTAINS iG'L>:
~U NI CD :N CR CA NA FE AL PB SI P
0.02 O-.^O O.OO J.OO 0.00 O.OO O.OO O.OO O.OO O.OO O.OO O.OO
•***««••>•$«•««<•«» ••$••••••*»«$••*•! •*••*••«!••••$«•»
t«(i «tt ?•!••« t««*«»»»a(»«*tt*<***t***««***»<*««*o»««»a*c*«*»»»
:; 'O THE STRIPPING PROCESS CONTAINS:
CU NI CO ZN CR CA NA FE AL PB SI P
5.ie .'.OO 0.00 O.OO 0.00 O.OO O.OO O.OO O.OO O.OO O.OO 0.20
THE PINAL AQUEOUS SOLUTION CONTAINS:
CU NI CO ZN CR CA NA FE AL PB SI P
0.02 3.22 O.OO 3.25 2.86 0.10 0.7Q 0.14 O.75 O.CO O.OO 0.50
• •t
-------
TABLE 8.147. CONTINUED
«M*fttMMMMt»t*M«**»SMMfitt«*tttttS*«t«««ttttaM3*««e«**S
SK ZINC AND IRON
• (••»MIMMMMt*»M««tII»M<)tI«S*M>*IM •>«>S«**»C**S9ttIIII>
IML SULVtNT EXTRACTION CONDITIONS ARE AS FOLLOWS:
ItMffRAUiRE (IN DEGREES CENTIGRADE)! SO
L'ONIACI I 1ME (IN MINUTES)! 3.O
O-A KAI 111: I. OH
IHL V'UIUHE F1OW KA'ttL/P) OF ORGANIC SOLUTION REQUIRED ISt 161716.
I HE VULUME % DfHPA IN THE ORGANIC SOLUTION IS: 40. OO
THE VOLUME % KERHAi' S1O IN THE ORGANIC SOLUTION ISl 60.OO
THE VOLUnt X 0 IN IHf. ORGANIC SOLUTION ISt O.OO
1VB.O StTS OF nikINU TANKS ARE REC'JIRED, 4 STAGES PER SET.
»M»lC«*«»IMIMM*»M»IO»M*teM*»«*«8««t****9S»««*tltt«*«tt>*
mi. riiitiK '..i.u;t i is i.zo
IHC txir^CTlON tTFUlENOS FOR STAGE 1 AKE:
•/ii Nl CD 2N CR CA NA FE AL PB SI P
O.Ot> %.OO .'-5.OO 55. OO O.OO 18.80 O.OO BO.00 27.00 O.OO O.OO 0.00
»•««*(»* *»*««**"«»»*« t**»*M*t*«»«*t»*«»*»8>*(t98*$*««»tt«««t*
1ME itOUEOllf- IA!T1NG ••TAGE 1 CONTAINS (G/L>:
i'U Nl CD :N CR CA NA FE AL PB SI P
o.->r :-.rr o.oo ..n 2.Bn o.oe o.'o 0.03 o.se o.oo «.i.oo o.so
•»••»•••*»»*t»0«»»»»»»«*»««•»«t»»«««»«»*»»«»»»«»8t*»*«»*»««*»«»
•HE ORGANIC FXIT1NQ STAGE 1 CONTAINS (G/L)I
rU Nl CD ZN CR CA NA FE AL PB SI F
O.O1 O.OO 0.00 3.19 O.OO 0.08 O.OO O.I4 O.61 O.OO O.OO 0.00
M«$««»»»«»t«»8»»»|tt«»»»*«««««»»»t»«»»»»•»«»«*«*•«»•«
»•««(fft**•»*««*•**»»l««*t•»••«*«**»*M*t***»««(S»*(t«**t««S«t<
THc PH FOR STAGE 2 IS =.00
THE EXTRACTION EFFICIENCVS FOR STAGE 2 ARE:
CU Nl CD ZN CR CA NA FE AL FB SI P
SO.OO 0.30 BS.OO 8E.OO O.OO 0.00 0.00 82.OO 59.00 O.OO O.OO O.IOO
»•(**««(»»««4«*f««*I•»«»•••*<>»**$*•««»«*«»**»«**«»<«**********
THE AQUEOUS EXITING STAGE 2 CONTAINS >G/L>I
CU Nl CD ZN CR CA NA FE AL PB SI P
O.M 3.22 O.OO 0.25 i.86 0.00 0.70 O.C1 O.24 O.OO O.OO 0.5O
i
487
-------
TABLE 8.1',?. CONTINUED
888888888818*888888888888888888888888888?8888888*88888888888888
THE URBAN1C EXITING STAGE 2 CONTAINS CG/L)I
CU NI CD ZN CR CA NA FE AL PB SI P
O.01 O.OO O.OO 2.OS O.OO O.O6 O.OO O.O3 O.44 O.OO O.OO O.OO
THE PH FOR STAGE 3 IS 1.5O
THE EXTRACTION EFFICIENCYS FOR STAGE 3 AREi
CU NI CO ZN CR CA NA FE AL PB SI P
30.OO O.OO 65.OO 65.00 O.OO 46.7O O.OO 55.OO 28.OO O.OO O.OO O.OO
*SS88888888888888888888888888«8»8/8*88888**88888888888888888*888
THE AQUEOUS EXITING STAGE 3 CONTAINS (G/L>»
CU NI CD ZN CR CA Nrt FE AL PB SI P
O.O1 3.22 O.OO O.09 2.86 O.05 0.70 O.OO O.17 O.OO O.OO O.5O
8888888888888888888888888888888888888888888888888*8888888888888
THE ORGANIC EXITING STAGE 3 CONTAINS (G/L)I
CU NI CO ZN CR CA NA FE AL PB SI P
O.OO O.OO O.OO 0.19 O.OO O.O6 O.OO O.OO O.10 O.OO O.OO O.OO
88888888888888888*88888888888888S888888*88*88*88C88<8*
88»«8»»*8M888888888888888888888*888»888888888888*888888888a»8*
THE PH FOR STAGE 4 IS 1.30
TH.= EXTRACTION EFFICIENCYS FOR STAGE 4 AREl
CU NI CD ZN CR CA NA FE AL PB 61 P
0.00 0.00 30.00 30.00 O.OO 41.40 O.OO 25.OO 1B.4O O.OO O.OO O.OO
88t*8888888888888888a8S888888»88*888888«8C88888888*88888888*38»
THE AQUEOUS EXITING STAGE 4 CONTAINS (6/L)l
CU NI CO ZN CR CA NA FE AL PB SI P
"...01 S.22 O.OO 0.06 2.86 0.03 0.70 0.00 0.14 O.OO O.OO O.5O
I,,•,**fM**l«*«*«888«»*888*88*8*88*8*8888888*88888888*8**8888*
t ORGHlilC £.) II ING STAGE 4 CONTAINS (G/L) I
CJ NI CO ZN CR CA NA FE AL PB SI P
O.OO O.OO O.OO O.03 O.OO 0.02 0.00 0.00 O.O3 O.OO O.OO O.OO
888888888888888888888888888888888888*8»8888888*88868«8
8888*88*8888*888888888888*88**888888888888888888888888t88*8**88
THE ORGANIC TO THE STRIPPING PROCESS CONTAINS*
CU NI CO ZN CR CA NA FE AL PB SI P
O.O1 O.OO O.OO 3.19 O.OO O.O8 O.OO O.14 O.61 O.OO O.OO O.OO
488
-------
TABLE 8.147. CONTINUED
THE FINAL -AQUEOUS SOUTH UN CONTAINS*
CU NI CD ZN CR CA MA FE AL PB SI P
0.01 3.22 O.OO O.06 2.86 0.03 0.7O O.OO O. 14 O.OO O.OO 0.5O
GENERAL EXTRACTION PROGRAM : CHROMII* OXIDATION
***«*»*«»*ti****»**«*sets*8«»t«**»*$«
GPL IN STARTING SOLUTION!
CU NI CD ZN CR CA NA FE AL PB SI P
O.O1 3.22 O.OO O.O6 2.86 O.O3 0.7O O.OO O.14 O.OO O.OO O.SO
THE LITERS/DAY STARTING SOLUTION! 161716.
THE PERCENT EXTRACTIONS:
CU NI CO ZN CR CA NA FE AL PB SI P
O.O O.O O.O O.O 14.7 O.O O.O 95.0 42.B O.O O.O O.O
POUNDS METALS ENTERING PROCESS:
CU NI CO ZN -CR CA NA TE AL PB SI P
2.6 114B.9 0.0 22.1 1019.4 9.5 24B.6 O.6 49.6 O.O O.O 179.O
NEW GPL VALUES - SOLUTION EXITING PROCESS i
CU NI CO ZN CR CA NA FE AL PB SI P
O.C1 3.22 0.00 0.06 2.44 O.O3 0.70 O.CJ O.OB O.OO O.OO O.SO
THE HOLE. UT. OF THE PRECIP. SPECIES:
CU
O.O
NI
0.0
THE MOLES OF
CU
l.O
THE PPD
CU
0.0
NI
1.0
CO
0.0
PRECIP.
CO
l.O
ZN
O.O
CR
103.0
CA
O.O
SPECIES PRODUCED
ZN
1.0
CR
l.O
CA
1.0
NA
0.0
FE
107.0
AL
7B.O
PB
O.O
SI
O.O
P
O.O
PER HOLE METAL i
NA
1.0
FE
l.O
AL
1.0
PB
l.O
SI
l.O
P
1.0
OF EACH PRECIP. SPECIESi
NI
O.O
CO
O.O
ZN
0.0
CR
296. 0
CA
O.O
NA
O.O
FE
1.1
AL
61.4
PB
O.O
SI
O.O
P
O.O
THE TOTAL PPD OF SOLID SPECIES PRECIP: 359.3
GPL ACID IN SOLNi O.O
489
-------
TABLE 8.147. CONTINUED
M8888888888*M8t88MMt8888888t»:88«
GENERAL EXTRACTION PROGRAM: LEAD CHROKATE PRECIPITATION.
8*8888*888888888888*88888888888888888
GPL IN START INS SOI.LlTIONi
CU NI CD 2N CR CA NA FE AL PB SI P
0.01 3.22 O.OO 0.06 2.44 O.O3 O.7O O.OO O.08 O.OO O.OO O.5O
THE LITERS/DAY STARTING SOL'JTIONi 161716.
THE PERCENT EXTRACTIONSI
CU NI CD ZN CR CA NA FE AL PB SI P
O.O O.O O.O O.O 99.6 C.O O.O O.O 100.0 O.O O.O O.O
POUNDS METALS ENTERING PROCESSl
CU NI CD 2N CR CA NA FE AL PB SI P
2.6 1148.9 0.0 22.1 869.6 9.5 248.6 0.0 28.4 O.O O.O 179.0
NEW GPL VALUES - SOLUTION EXITING PROCESS I
CU NI CD ZN CR CA NA FE AL PB SI P
O.01 3.22 O.OO O.06 O.01 O.O3 O.7O O.OO O.OO O.OO O.OO O.SO
THE HOLE. WT. OF THE PRECIP. SPECIES*
CU NI CD ZN CR CA NA FE AL PB SI P
O.O O.O O.O O.O 323.1 O.O O.O O.O O.O O.O O.O O.O
THE MOLES OF PRECIP. SPECIES PRODUCED PER HOLE METALI
CU NI CD ZN CR CA NA FE AL PB 81 P
l.O l.O 1.0 l.O l.O l.O l.O l.O l.O 1.0 l.O l.O
THE PPO OF EACH PRECIP. SPECIESI
CU NI CD ZN CR CA NA FE AL PB SI P
O.O O.O O.O O.O 9380.3 O.O O.O O.O O.O O.O O.O O.O
THE TOTAL PPD OF SOLID SPECIES PRECIP! 5380.3
GPL ACID IN SOLNi O.O
****8*8*888888**M*88*8*88t8888*M88*
GENERAL EXTRACTION PROGRAM : NICKEL SULFIDE WECIPITATION.
88888C8884888888888888888888S8888888*
490
-------
TABI.E 8.147. CONTINUED
GPL I* STARTING BOLUTIONi
CU NI CO 2N CR CA NA KE AL PB 81 P,
O.O1 3.22 O.OO O.06 0.01 O.O3 O.7O O.OO O.OO O.OO O.OO O.SO
THE LITERS/DAY STARTING SOLUTIONi 161716.
THE PERCENT EXTRACTIONS!
CU NI CO ZN CR CA NA FE AL P6 SI P
9S.O 99.6 95.O 95.0 10.0 O.O O.O 9S.O O.O 9S.O O.O 1OO.O
POUNDS fCTALS ENTERING PROCESSI
CU NI CO ZN CR CA NA FE AL PB 81 P
2.6 IMG.9 0.0 22.1 3.5 9.5 248.6 O.O O.O O.O O.O 179.O
NEW GPL VALUES - SOLUTION EXITING PROCESS I
CU NI CD ZN CR CA NA FE AL PB 61 P
O.OO O.01 O.OO 0.00 0.01 O.03 O.7O O.OO O.OO O.OO O.OO 0.00
THE MOLE. UT. OF THE PRECIP. SPECIESl
CU NI CD ZN CR CA NA FE AL PB 81 P
96.0 91.0 144.0 97.0 0.0 O.O O.O 88.0 O.O 239.O O.O 9S.O
THE HOLES OF PRECIP. SPECIES PRODUCED PER PSXE KETALl
CU NI CD ZN CR CA NA FE AL PB SI P
l.O l.O 1.0 1.0 1.0 l.O l.O 1.0 1.0 1.0 I.O 1.0
••HIT FFD OF t>.rH PRECIP. SPECIESl
. U HI :D ZN CR CA NA FE AL PB SI P
'..- 1777.; <>.••> 31.2 0.0 0.0 O.O O.O O.O O.O O.O 948.8
IHi TOTAL PFD OF SOLID SPECIES PRECIPl 2361.1
CPL ACID IN SOLNi O.O
491
-------